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With numerous Plates, Maps, and Illustrations. 21s. net. 


A Practical Treatise on the Cyanide Process ; its Application, Methods 
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*** This book deals with the Cyanide Process from Technical, Commercial, and Scientific 
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A Three Years' Course for Students of Schools of Mines. 

By W. A. MACLEOD, B.A., B.Sc, A.O.S.M. (N.Z.), Director, Government School 
of Mines, Charters Towers, Queensland; and CHAS. WALKER, F.C.S., Lecturer in 
Chemistry and Metallurgy, Government School of Mines, Charters Towers, Queensland. 

London : CHARLES GRIFFIN & CO., LTD., Exeter Street, Strand, W.C. 

[" The Cyanide Process of Gold 

[Published by CimiLKa Griffin & Co., Ltd., London .] 














[The First English Edition was revised and enlarged from the Third Edition 

published in New Zealand.] 

itl) JTronttgpieee, opiate* anb Illtxgt ration*. 

(Authorized Text-book. Australian Schools of Mines.) 



[All Rights Reserved.] 

A Complete Catalogue of Works on Engineering, 
Mining, and Metallurgy will be forwarded 
post free on application to the publishers. 

78161 ' <o**'bOA 

APR 14 1904 


•Pa. i 


The favourable reception accorded to earlier editions has en- 
couraged me to revise the old matter, and at the same time add 
much new material, which, for the most part, relates to lead- 
smelting of gold-slimes, the treatment of sulpho-telluride ores, and 
filter-press practice. In Western Australia, the adoption of filter- 
pressing was mainly determined by a combination of peculiar 
local conditions, namely, the natural tendency of the gold ores and 
matrix to form slimes, the scarcity of fresh water, and the saline 
character of the only water available for milling purposes. The 
density of the brackish waters rendered the settlement of the 
finer material in the decantation process so slow, laborious, and 
imperfect as to make the use of filter-presses almost imperative 
for quick and effective treatment. The slimes, it should be noted, 
formed the first or primary product of the mills, and were of high 
value. In South Africa, where the question of filter-pressing 
versus decantation has been engaging attention for some time, 
the slimes are in all cases a secondary product and of low-grade 

Filter-pressing has been adopted with much success by the 
Waihi Company, in New Zealand, for the treatment of ordinary 
slimes, but, on the other hand, it should be mentioned that several 
neighbouring companies are satisfied with the decantation process 
for the treatment of similar material. It is quite clear that the 
relative merits of the two processes must, in every case, be deter- 
mined by exhaustive trials extending over a period of, say, four to 


six mouths, so as to eliminate the element of fortuitous chance 
in favour of either, and thereby enable a reliable estimate of 
costs to be prepared. With South African and New Zealand 
conditions, the solution of the problem is obviously one of cost. 
In Western Australia, the question of cost is subordinate to 
that of expediency. 

The introduction of lead-smelting of gold-slimes marks a 
notable advance in cyanide practice, and if the claims of Mr. 
Tavenor, the author of the process, are verified by more com- 
prehensive trials — as indeed seems most probable — lead-smelting 
will in a short time displace the old smelting and acid- treatment 
processes, at any rate in all the larger permanent cyanide plants. 

There is still much room for special chemical investigation in 
several directions, and in this respect the researches of a number 
of accomplished metallurgical chemists in South Africa have 
added much valuable material to the literature of the cyanide 
process. The successful regeneration of foul cyanide solutions is 
still unsolved, and is at present the subject of investigation by 
several American chemists. The results of their work will be 
awaited with much interest. 

I have to acknowledge my indebtedness to Mr E. G. Banks, 
M.I.M.M., and Mr Frank B. Allen, M.A., B.Sc, for special 
contributions on Waihi slime and filter-press practice and Kal- 
goorlie sulpho-telluride practice respectively; and to the pro- 
prietors of the Mining Journal, Mining and Engineering Journal, 
and Australian Mining Standard for permission to make extracts 
from articles which appeared in their columns at different times. 
In these and all other cases, due acknowledgment and reference 

are made in the text. 


University, Dunedin, N.Z., 
September 1903. 



The favourable reception accorded to the " Third Edition of this 
work, published in New Zealand, has enabled the author to again 
revise and enlarge the text, with a view of placing before his 
readers, earlier than was anticipated, the latest information 
available on this progressive branch of metallurgy. The general 
plan of the original work, which was intended for the use of 
mining students, metallurgists, and cyanide operators, has been 
retained in the present edition, which is the first published in 
England, and into which only such changes have been introduced 
as the author considered needful. Of late years the application 
of scientific investigations and methods to the treatment of ores 
has rendered metallurgy more and more dependent on chemical 
knowledge, and in no department is this more obvious than in 
the Cyanide Process of Gold Extraction, which often presents 
the most perplexing problems, due, in many cases, to the simplest 

In the present edition several new illustrations and tables have 
been added, while the information relating to the treatment of 
slimes, the analyses of solutions, and cyanide poisoning, has been 
greatly extended ; while, by the adoption of slightly smaller type, 
and closer setting, the actual number of pages has been reduced, 
although the text has been considerably enlarged. 

Since the revision of these pages, wet-crushing and cyanide 
treatment have largely superseded dry-crushing and direct 


cyaniding in New Zealand, and in every case their installation has 
been attended with complete success, notwithstanding the large 
proportion of slimes at some of the mines. The ores are mostly 
chalcedonic in character, and contain both coarse and fine gold 
associated with silver sulphide. The silver occurs in constantly 
varying proportions, requiring varying strengths of solution to 
obtain adequate extractions, and unremitting care on the part of 
the metallurgist in charge. 

The results obtained in the treatment of these comparatively 
complex ores are of world-wide interest, and have been embodied 
by the author in a separate appendix. 


Auckland, January 1900. 




The M c Arthur Forrest- Process, 1 

Chemistry op the Process, 4 

Laboratory Experiments, 17 

Control, Testing, and Analysis of Solutions, .... 22 

The Appliances and Plant, 48 

The Actual Extraction by Cyanide, 69 

The Treatment of Slimes, 80 

Treatment of Concentrates, 98 


Leaching by Agitation, • • 96> 



CHAPTER X. page 

Zinc Precipitation and Treatment op Gold Slimes, . . .100 

The Application of the Process, 121 

The Sibmens-Halske Process 171 

Other Cyanide Processeb, 177 

Antidotes for Cyanide Poisoning, 184 

Index, 188 




Mammoth Cyanide Plant, S. Africa, . . . Frontispiece 

I. Wooden Leaching Vat and Supports, .... to/ace 52 

II. Steel Vat, „ 54 

II a. Park's Improved Bottom- Discharge Door, . . ,, 58 

III. Side-Discharge Door, ,,60 

IV. Zinc Extractor Box, . . . . . . . ,, 62 

IVa. Zinc Extractor Box, New Pattern, . . . . ,, 64 

V. Butters' Distributor, and Roasting Furnace, . ,, 74 

VI. Butters' Distributor, Plan and Elevation, . ,, 74 

VII. Tailings Plant ,,96 

VIII. Precipitation Room, Waihi, ,,136 

IX. Slime Agitators at Waikino, . . . ,, 138 

IX. Roche's Bottom-Discharge Door,. . . . ,, 138 


Fig. 1. Showing Turn-Buckle 53 

Fig. 2. Butters' Bottom-Discharge Door, 57 

Fig. 3. Irvine's Bottom-Discharge Door, 58 

Fig. 4. Side-Discharge Door, . . . . . .59 







It has long been known that gold and silver are soluble in 
solutions of alkaline cyanides, but it is only within the past few 
years that this knowledge has been applied on a commercial scale 
to the extraction of the preoious metals from their ores. 

The discovery of the fact that the dilute solution of potassium 
cyanide is a solvent for natural gold, ranks among the most 
remarkable discoveries of the present century in metallurgical 
science ; and the widespread and successful application of the fact 
must mark an epoch in the history of gold extraction for all time. 


The cyanide process can be applied with success to the treat- 
ment of free-milling ores in which the gold occurs in fine particles, 
or of tailings and concentrates resulting from wet-crushing and 
copper-plate amalgamation, or dry-crushing and pan-amalgama- 
tion. It can also be used for the treatment of many so-called 
refractory ores, especially those in which the gold occurs in such 
a finely-divided form that even amalgamation in pans fails to 



recover a satisfactory percentage of the values; or of ores in 
which the gold is coated with a film of metallic oxide or sulphide, 
rendering it non-amalgamable, and ores in which the gold is asso- 
ciated with, or entangled in, a highly pyritic matrix. 

All the common ores of silver are more or less soluble in dilute 
solutions of cyanide. Those most readily soluble are the chloride 
(AgCl) and the sub-sulphide (Ag 2 S), and these are fortunately the 
most abundant ; but the rate of dissolution of silver and its ores 
is much slower than that of gold, and is accompanied by a higher 
consumption of cyanide. 


The cyanide process cannot be applied with success to the 
treatment of ores in which the gold occurs even in a fairly coarse 
condition. When an ore contains a proportion of both fine and 
coarse gold the cyanide process may be used to extract the fine 
gold, but a supplementary treatment will have to be used to 
recover the coarse gold, since the slowness of the dissolution 
would take too long for a commercial basis of working. 

With free-milling ores of the latter class the recovery of the 
coarse gold is generally effected by copper-plate amalgamation ; 
and, in the case of wet-crushing, this treatment precedes the 
cyanide leaching, while, in the case of dry-crushing, it follows it. 

The experience gained during the use of the cyanide process has 
shown that solutions of potassium cyanide, even when very dilute, 
act most energetically on all the sulphide, oxide, and carbonate ores 
of copper, and also on the sulphides of antimony and bismuth ; 
hence, when any of these is present, even in small proportion, 
the treatment of the ore becomes difficult, and sometimes impos- 
sible, on account of the great consumption of cyanide. In practice 
it is found that an unduly large consumption of cyanide is generally 
accompanied by a low rate of extraction of the gold and silver 
contained in the ore or tailings. 

From the foregoing it is obvious that the process will be most 
successful in the treatment of ores in which the gold occurs in a 
very fine state, and in which the quantity of base minerals or 
metallic salts, destructive to cyanide, is small. 

Further, the author ascertained as the result of many simul- 
taneous working trials in the N.Z. Government Metallurgical 
Works that argentiferous gold ores which were amenable to treat- 
ment by the Washoe pan-amalgamation process, in most cases 
yielded better results by cyanide treatment, even when they con- 
tained a small percentage of zinc and lead. 


An intelligent knowledge of first principles backed by experience 
of working details and working requirements has led to many 
ingenious adaptations, the tendency of which has been to greatly 
widen the scope of the original cyanide process. A notable case 
in point is the successful treatment of the rich sulpho-telluride 
ores of America and Kalgoorlie, which affords satisfactory evidence 
of the progressive trend of this important branch of metallurgy. 



When gold is acted on by an aqueous solution of potassium 
cyanide, a solution is obtained which, when evaporated, yields 
octahedral crystals having the composition of the auro-potassic 
cyanide (AuKCy 2 ), which is a double cyanide of gold and 

The exact reaction which takes place when gold is dissolved by 
potassium cyanide is not yet well understood, being still a subject 
of much doubt and uncertainty. According to some authorities, 
the gold is oxidized before it is dissolved ; while others maintain 
that the cyanide is first oxidized and then acts on the gold. 

The reaction suggested by Eisner in 1842 is the one now most 
generally accepted by chemists. It is represented by the follow- 
ing equation : — 

4Au + 8KCy + 2 + 2H 2 = 4 AuKCy 2 + 4KHO. 

According to the above equation, an ounce of oxygen is 
required for every pound of potassium cyanide employed for the 
dissolution of the gold. This view has received substantial sup- 
port from the author's experiments in 1891, and from those of 
Skey* in 1892, and has since been proved experimentally correct 
by Maclaurin t in his classical paper on the subject 

The valuable researches of Skey and Maclaurin have shown that 
the rate of dissolution of pure gold, under theoretical conditions, 
reaches a maximum in passing from dilute to concentrated solu- 
tions of potassium cyanide. By actual experiment, it was proved 
that the maximum rate was reached with a 0*25 per cent, solution 
of cyanide. On a working scale the maximum varies with the 
character of the mineral constituents of the ore, and can easily be 
determined by a series of laboratory experiments. 

A weak solution is always more active than a strong one, and 

* Skey, N.Z. Mines Report, 1894. 
f Jour. Chem, Soc., May 1893, p. 724. 


Maclaurin considers that this remarkable fact may be accounted 
for by supposing that the rate of dissolution of gold is partly 
dependent on the number of cyanide molecules in a unit volume ; 
and partly on the number of oxygen molecules in the same 
volume. One of the most important results of his exhaustive 
experiments was the demonstration of the fact that the solubility 
of oxygen in cyanide solutions decreases with concentration of the 

Weak aqueous solutions of cyanide exert a very marked 
action on gold and silver when these metals are associated with 
ores of copper and antimony. This circumstance becomes very 
prominent during the treatment of cupriferous ores on a large 

The cyanides of the alkaline metals are soluble in water, while 
those of the heavy metals, with the exception of gold and mercury, 
are insoluble. The insoluble salts are, however, soluble in excess 
of potassium cyanide. 

The use of an oxidizing agent that will readily part with a 
portion of its oxygen in a cyanide solution forms the essential 
feature of several new patent cyanide processes. The employ- 
ment of such an agent serves to accelerate the dissolution of the 
gold contained in the ore. The artificial aeration of cyanide 
solutions is undertaken to supply atmospheric oxygen with the 
same object. 

Consumption of Cyanide. — According to Eisner's equation, 
about 4*5 lbs. of cyanide should dissolve 100 ounces of gold, but 
in practice it is found that it takes nearly forty times that 
quantity. The causes which operate in the practice of the process 
to effect so large a consumption of cyanide, over that required by 
Eisner's simple equation, are at present not fully investigated. 

To dissolve 100 oz. of silver would require 7*5 lbs. of cyanide, 
according to the equation : — 

4Ag + 8KCy + 2 + 2H 2 = 4(AgKCy 2 ) + 4KHO. 

For the dissolution of 100 oz. of silver existing as the sub- 
sulphide (Ag 2 S), 7*01 lbs. of cyanide would be required by the 
following equation': — 

Ag 2 S + 4KCy = 2(AgKCy 2 ) + K 2 S. 

The potassium sulphide resulting from the dissolution of silver 
sulphide also tends to cause a further loss of cyanide by precipi- 
tating gold which will require an excess of free cyanide to 
redissolve it. It is the need for this excess of cyanide which 
necessitates the use of ^comparatively strong solutions in the 
treatment of argentiferous gold ores. 


Potassium cyanide is, chemically, a most active organic com- 
pound, possessing the property of forming so large a number of 
complicated and unexpected combinations in the presence of 
mineral acids and base metals, that its reactions and behaviour 
with different classes of ore, and under varying conditions, can 
only be unravelled by much patient research, both in the labora- 
tory and under working conditions. 

During the treatment of ores by the cyanide process, the most 
puzzling difficulties are continually met with, requiring the con- 
stant care and attention of the metallurgist in charge. 

Causes of LOSS of Cyanide. — Some of the principal and 
more obvious causes of the enormous loss of cyanide which takes 
place in the working of the process are as follows : — 
~ 1. Loss by absorption in wooden vats or tanks. 

- 2. Loss by decomposition by atmospheric carbon dioxide. 

3. Mechanical loss in residues, and by dilution of solutions 
during washing. 

- 4. Loss by decomposition due to the presence of mineral acids 

and salts. 

5. Loss due to presence of ores soluble in cyanide. 

6. Loss when gold exists as amalgam. 

7. Loss due to the presence of charcoal in kiln-dried ore. 
Loss by Absorption in Vats. — This is especially noticeable 

in new plants. At the Witwatersrand Goldfields, the loss from 
this cause is said by Mr. C. Butters to amount to a pound of 
cyanide per ton of tailings treated. At the first monthly " clean 
up " in a new plant, the actual extraction is often twenty or more 
per cent, below the theoretical, but after a few months it gener- 
ally rises to within three to six per cent, of the extraction as 
determined by assay. 

With iron or steel vats there is no appreciable loss by absorption. 

Loss due to Decomposition by Atmospheric Carbon 

Dioxide. — The carbonic acid gas of the atmosphere decomposes 
potassium cyanide with the formation of potassium carbonate, 
and the liberation of hydrocyanic (prussic) acid, thus : — 

2KCy + C0 2 + H 2 = K 2 C0 3 + 2HCy. 

The prussic acid thus liberated would be neutralized by any 
caustic alkali present in the cyanide solution. 

Mechanical Loss in Residues, and by Dilution during 

'Washing. — During washing there is an inability to extract the 
whole of the cyanide from the residual tailings. The dilution of 
the cyanide solutions also occasions a loss of cyanide in washing. 
A large quantity of dilute cyanide solution is formed, a portion 
only of which can be utilized to make up fresh solutions. 


Loss by Decomposition due to Mineral Acids and 

Salts.— The metallic minerals most commonly found associated 
with gold in quartz veins are iron pyrites, copper pyrites, zinc 
blende, galena, and antimouite. By far the most common and 
abundant of these is iron pyrites. 

It has been shown by Skey and others that clean fresh iron 
pyrites is not acted on by working solutions of cyanide. The 
decomposition products of this mineral, however, act most 
destructively on cyanide, and the obvious conclusion to be drawn 
from this is that the treatment of pyritic tailings, or concentrates, 
by cyanide should be undertaken with as little delay as possible ; 
more especially when the pyrites occurs in the marcasite form, 
which is much more prone to oxidation than the cubical or 
isometric form. 

In the shallow parts of mines, the pyrites is generally oxidized 
to ferric oxide, which does not act chemically on cyanide, but 
causes a mechanical loss through the formation, both in wet and 
dry crushing, of extremely fine slimes, which are very absorbent 
and retentive of cyanide solutions. 

Iron pyrites (FeS 2 ) is decomposed by atmospheric oxygen in 
the presence of moisture into the soluble ferrous sulphate and free 
sulphuric acid, according to the following equation : — 

FeS 2 + H 2 + 70 = FeS0 4 + H 2 S0 4 . 

In the kiln-drying of ores to be dry-crushed, the heat to which 
the ore, often in large pieces, is subjected, is not very uniform, 
especially in large kilns. With pyritic ores the sulphides are 
decomposed at certain temperatures into oxides and soluble sul- 
phates ; and at higher temperatures the latter salts are converted 
into oxides. 

The steam generated from the moisture in the fuel and in the 
ore itself assists these reactions. In the kiln, where the tempera- 
ture is high, reducing gases are evolved, and these may impede 
the oxidation of the sulphide, causing the formation of lower sul- 
phides and basic sulphates, which are insoluble in water, but react 
on cyanide. 

For these reasons, and also for economy, ores intended for dry 
treatment by cyanide should never be kiln-dried. In the case of 
developed mines, with an assured and steady output, there is 
nothing to justify the practice, but much to condemn it. 

As we have seen, the atmospheric oxidation of pyrites results in 
the production of free sulphuric acid and ferrous sulphate. This 
ferrous sulphate may in turn be decomposed, by the action of the 
air, into insoluble basic sulphates. Thus, partly oxidized pyritic 
ores or tailings may contain free sulphuric acid, soluble ferrous 


sulphate, insoluble basic sulphates, and probably also traces of 
other basic salts of complex and variable composition, all of which 
react upon solutions of potassium cyanide, thereby causing a loss 
of cyanide. 

The reactions which are most likely to take place in acid ores 
or tailings in the presence of cyanide are : — 

(a.) The liberation of hydrocyanic acid. 

(b.) The formation of ferro- and ferri-cyanides. 

The free acids in the ore react on the cyanide as shown by the 
equation : — 

2KCy + H 2 S0 4 = 2HCy + K 2 S0 4 . 

Feldtmann considers it possible for the hydrocyanic acid thus 
liberated to diffuse itself through the ore and dissolve appreciable 
quantities of gold.* For this reason he strongly condemns the 
practice of washing acid tailings in the leaching vats, as these must 
always contain a residual portion of cyanide from which prussic 
acid would be liberated. Any gold dissolved by this gas would 
be carried away in the water or alkaline wash ; and to avoid this 
possible source of loss, which he thinks may account for the 
mysterious discrepancy sometimes found between the assay and 
the actual extraction, he recommends the system of washing in one 
vat and leaching in another. On the other hand, Skey, when dis- 
cussing this subject with the author, stated that hydrocyanic acid 
was not a solvent for gold. It is obvious, however, that this 
liberated HCy in the presence of any residual alkali in the vat, . 
would form an alkaline cyanide capable of dissolving and removing 
gold, and Feldtmann probably had this combination in his mind 
at the time of writing. 

Of the iron salts, the one of most common occurrence in pyritic 
ores or tailings is the soluble ferrous sulphate (FeS0 4 ), which 
reacts with potassium cyanide to form potassium . ferro-cyanide 
and sulphate, thus : — I 

FeS0 4 + 6KCy = K 4 FeCy 6 + K 2 S0 4 . 

The potassium ferro-cyanide thus formed is, in its turn, reacted 
on by any excess of ferrous sulphate still present with the produc- 
tion of Prussian blue according to the equation : — 

3K 4 FeCy 6 + 6FeS0 4 + 30 = Fe 2 3 + 6K 2 S0 4 + Fe 4 (FeCy 6 ) 3 . 

A blue colour in the solution, on the surface of the tailings, or 
in the seams of the staves of the vats, indicates a large consump- 

* Feldtmann, Notes on Gold Extraction, p. 5. 




tion and loss of cyanide due to imperfect washing and neutraliza- 
tion of the acidity in the preliminary treatment. 

A white soum or precipitate is sometimes seen on the surface of 
the solutions when they are coming off acid. This precipitate 
turns into Prussian blue by exposure to the air and light. 

The normal ferric sulphate Fe 2 (S0 4 ), is insoluble in water, 
and oannot be removed by ordinary water-washing. It reacts 
with potassium cyanide, causing a loss of cyanogen due to the 
liberation of prussic acid and the formation of the ferric hydrate, 
as shown by the two following equations : — 

Fe 2 (S0 4 ) 8 + 6KCy « Fe 2 Cy 6 + 3K 2 S0 4 

Fe 2 Cy 6 + 6H 2 = Fe 2 (HO) 6 + 6HCy. 

It is probable that in most partially oxidized pyritic ores and 
tailings the ferrous and ferrio sulphates exist together, the former 
in large excess. In this case the decomposition of the cyanide 
would result in the production of ferrous cyanide and potassium 
sulphate, thus : — 

12KCy + 3FeS0 4 + Fe 2 (S0 4 ) 3 = Fe 8 (FeCy 6 ) 2 + 6K 2 S0 4 . 

In the case of earthy pyritic ores, the weathering or oxidation of 
the metallic sulphides would result in the production of sulphates 
of magnesia, lime, or alumina. The action of these sulphates is 
not very clear, but they most likely react on cyanide with the 
liberation of prussic acid, accompanied by the formation of the 
hydrated oxide of the metals in question, which would be pre- 
cipitated as an insoluble incrustation in the solution pipes. 

The above reactions clearly emphasize the necessity of a most 
careful preliminary alkaline treatment of pyritic material, in order 
to avoid undue loss of cyanide, and ensure satisfactory results. 

All the iron salts and earthy sulphates can be rendered innocu- 
ous by the application of an alkali before treatment with the 
oyanide. By this means all the soluble iron salts are precipitated 
as ferrous hydrate, which rapidly oxidizes to ferric hydrate ; while 
the basic ones soon oxidize in the presence of the alkali. It is 
important to remember that the alkali should be applied before, 
and not with the cyanide solutions, as these iron salts will destroy 
the cyanide as much in a strongly alkaline as in a nearly neutral 
solution. When the tailings contain free acid only, alkali and 
cyanide should be applied together. 

Loss due to Presence of Ores Soluble in Cyanide.— 

The sulphide, oxide, and carbonate ores of copper, and the sul- 
phides of antimony and bismuth, are acted on by potassium 


cyanide both in weak and strong solutions, and thereby cause a 
loss of cyanide in proportion to their abundance in the ore. In 
the treatment of an ore containing as little as 0*25 per cent, of 
copper the consumption of cyanide will be doubled. 

It is during the treatment of cupriferous ores that the selective 
action of weak cyanide solutions becomes most apparent. An ore 
may contain sufficient copper to decompose a 1 per cent, solution 
of cyanide and give a low extraction of gold, whereas a 0*35 per 
cent, solution would dissolve proportionately less copper, and give 
a fairly satisfactory extraction of the gold. But the same results 
would be obtained even in the absence of copper, for it has already 
been shown that the rate of dissolution of gold reaches a maxi- 
mum in passing from dilute to strong solutions. Hence a 0*35 
per cent, solution should extract more gold than a 1 per cent, 
solution, the weaker solution being nearer the strength at which 
a maximum rate of dissolution occurs, and which has been proved 
experimentally to be a 0*25 per cent, on pure gold. 

The cyanide treatment of ores, and zinc»precipitation of the 
gold, have shown the existence of copper in ores in which no trace 
of that metal could be detected, even by the most rigid chemical 
examination on large samples. An instance of this came under 
the notice of the author, at the Crown mines at Karangahake. 
The ore being treated there consisted of almost pure white quartz, 
free from all metallic impurities ; nevertheless, a portion of the 
zinc in the precipitation boxes was often coated with a film of 
bright metallic copper. The copper could not be derived from an 
outside source, or from any of the mechanical fittings in the mill 
or cyanide plant, and Mr. James Napier, the metallurgist and 
chemist in charge, was of the opinion that it existed in the ore 
in an infinitesimally small quantity, and only became manifest 
on the zinc turnings after the treatment of hundreds of tons of 

Copper pyrites is oxidized to the soluble sulphate at low tem- 
peratures, and this salt requires a greater heat to decompose it 
than iron pyrites. It is, therefore, probable that a portion, at 
least, of this mineral present in an ore, being dried in kilns, pre- 
paratory to dry-crushing and direct cyanide treatment, would be 
sulphatized, and thereby cause an appreciable loss of cyanide in 
a manner similar to that caused by the decomposition products of 
iron pyrites. 

Malachite and azurite, the green and blue carbonates of copper, 
are both readily soluble in dilute solutions of cyanide, with the 
production of copper-potassic cyanide and liberation of prussic 

Antimonite, the grey sesqui-sulphide of antimony,' is also readily 


acted on by weak cyanide solutions. It is frequently met with 
in the gold-bearing oars of the Thames and Beef ton goldnelds. 
The presence of a small percentage of antimonite in the large 
accumulation of tailings at Boatman's Creek, near Reefton, is 
said to have caused all attempts to treat them to end in failure, 
chiefly owing to the large consumption of cyanide and the low 
rate of extraction. 

Loss of Cyanide when Gold Exists as Amalgam. —It 

is well known to most millmen that a considerable portion of the 
gold in tailings, resulting from copper-plate amalgamation or pan- 
amalgamation, exists in the form of amalgam. When such 
tailings have to be treated the cyanide has to dissolve the 
mercury as well as the gold, thus causing a larger consumption 
of the solvent than would be necessary if the gold existed in a 
free state. 

According to Gmelin, mercury is not dissolved or acted on by 
potassium cyanide; but the practical working of the cyanide 
process has shown that his conclusion is contrary to actual 

At the cyanide works of the Cassel Gold Extracting Company, 
at Waihi, where a large stack of tailings and residues from pan- 
amalgamation were treated, 75 lbs. of mercury were collected in 
the condenser attached to the furnace for roasting the zinc slimes. 
The mercury thus recovered was only a small prpportion of the 
mercury dissolved by the cyanide, and afterwards precipitated in 
the zinc-extractor with the bullion. On every occasion when the 
roasting of the zinc slimes was being conducted, so much mercury 
was volatilized that the vapours pervaded every part of the 
buildings, condensing on every cool surface, and amalgamating 
all objects of gold and silver worn by the workman. 

The volatilization of mercury during the roasting of zinc slimes, 
resulting from the treatment of tailings, was noted by the 
author on several occasions at the Government Experimental 
Cyanide Works, and is of frequent occurrence at the cyanide plants 
at Kuaotunu. The same circumstance was noted by Dr. Scheidel 
at the Sylvia Cyanide Works at the Thames, where mercury was 
found in the zinc-bullion in considerable quantities. 

The mercury generally occurs in the tailings in the form of 
• amalgam in a very fine state of subdivision, and is dissolved by 
the cyanide, together with the associated gold and silver. It is 
precipitated with the bullion in the zinc precipitation boxes. 
When- the zinc slimes are oxidized the greater portion of the 
mercury is volatilized. 

Loss of Cyanide due to the Presence of Charcoal in 

the Ore. — It has long been known to chemists that charcoal 


cyanide will necessarily occur ; but with these sulphides, in addi- 
tion to loss of cyanide, there will be a loss of gold, and a still 
greater loss of silver, in proportion to the quantities present. 
This loss is brought about by the sulphur — that is, the alkaline 
sulphide — sulphurizing these metals to form sulphides with them,* 
the sulphide film so formed upon the metal preventing, or greatly 
retarding, the proper action of the cyanide solution. 

That gold does combine, and very readily, with the sulphur of 
both the alkaline sulphide and of hydrogen sulphide, Skey has 
already shown, t 

It is to precipitate the sulphur that gets into the cyanide in the 
cyanide process that Mr. M c Arthur has proposed to use, or does 
use (as per patent), a soluble lead salt dissolved in the cyanide. 

The problem for the chemist at the cyanide works is to find a 
practical method, whereby all the sulphur of antimonial and cup- 
rous sulphides can be made to combine with the cyanogen, rather 
than with the potassium of the cyanide. 

. The following results obtained by Skey show how extremely 
objectionable alkaline sulphides are, when present in the cyanide 

A rather strong solution of the cyanide, containing a small 
proportion of sulphur, was placed over a strip of gold coupled with 
a piece of copper-glance (sulphide of copper), but no solution of 
gold was perceived ; however, on substituting chalcopyrites for the 
copper-glance, the gold was rapidly removed. 

This experiment shows that the gold was sulphurized at the 
outset by the alkaline sulphide present in the cyanide, and that 
it required connecting with a substance of a strongly negative 
kind in order to effect the decomposition of the auriferous sul- 
phide so formed. 

Further experiments of a different kind showed that while pure 
1 per cent, cyanide solution dissolved a given weight of gold in 
ten minutes, a solution of the salt of the same strength, but con- 
taining T ^ part of sulphur (as a ^sulphide), % required two 
hours to dissolve the same weight of gold. The speeds were as 
12 to 1 in favour of the pure cyanide. 

The following results show to what extent even a gentle sul- 
phurizing, or flouring of the gold, interferes with its solution : — 

Gold sulphurized 60 seconds in K 2 S, dissolved in cyanide in 
62 minutes. 

* Trans. N.Z. Inst., vol. iii p. 216. 
t Trans. N.Z. Inst., vol. xxi., 1888. 

% Trans. N.Z. Inst., vol. xxi, 1888, "On the Preparation of Artificial 


Gold sulphurized 54 seconds in K 2 S, dissolved in cyanide in 50 

Gold sulphurized 1 second in K 2 S dissolved in cyanide in 36 

Gold, clean, dissolved in cyanide in 12 minutes. 

The gold was well washed from adherent potassic sulphide 
before being placed in the cyanide. Making clean gold the unit, 
the approximate times of dissolution are 1:3:4:5. 

Action of Sulpho-cyanides. — It has been held by some 
metallurgists that the presence of these in working solutions of 
cyanide is injurious, retarding the dissolution of the gold. As the 
result of much research, Godfrey Doveton, formerly of Camp Bird 
Mills, Ouray, Colorado, informs me that he has found that a solu- 
tion containing potassium sulpho-cyanide up to a certain point 
was more active than a solution of corresponding strength in 
KCy, with sulpho-cyanide present, and that even 2-50 grams, 
of KCyS in 100 c.c. of solution did not influence the extraction 
unfavourably. The salt alone, in solution in water, is a slow 
solvent for gold. 

Experiments on mill solutions show much the same result, and 
he has come to the conclusion that the presence of even consider- 
able quantities of sulpho salt in working solutions should not 
cause any uneasiness. 

The Action of Manganese Oxides on Cyanide.— 

During the treatment of a parcel of ore from the Komata gold-mine, 
near Waitekauri, the author fouud there was an unusual consump- 
tion of cyanide. The ore consisted of soft mullqcky, friable 
quartz, coloured quite black by a large percentage of pyrolusite^ 
and wad, and containing a trace of nickel and cobalt. 

A series of experiments were afterwards made to determine the 
cause of the loss, and the results of these at first led the author 
to the conclusion that the manganese oxides oxidized a portion of 
the cyanide to cyanate* It is well known that pyrolusite parts 
with a portion of its oxygen under the influence of heat alone, but 
more readily so in the presence of an easily oxidizable substance. 
Further research showed, however, that the loss was due to the 
cobalt in the ore, which dissolved somewhat readily, thus consum- 
ing cyanide. It is interesting to note that the dissolved cobalt 
was precipitated with the gold and silver on the zinc in the ex- 
tractor box, and, like copper, was found to interfere with the 
precipitation of the gold. 

The Action of Oxygen-bearing Agents.— It seems probable 
that in ores containing copper or other base metal soluble in a 
solution of potassium cyanide, the base metal would, from its pre- 
ponderance in the ore, necessarily utilize the greater amount of 


the available oxygen, thereby tending to render the dissolution of 
the gold slow and imperfect. Hence it is reasonable to conclude 
that the employment of an oxidizing agent that would supply the 
deficiency of oxygen in such complex gold containing ores would 
be beneficial. That the use of an oxygen-bearing agent is justi- 
fied in special cases, seems well established by* the experience of 
the author and the results reported by reliable metallurgical 
chemists in South Africa, America and Western Australia. 



The cyanide process is essentially a chemical one, and a commodi- 
ous and well-equipped laboratory forms one of the most important 
and necessary parts of the whole plant. 

It is the duty of the metallurgist in charge to determine, by 
actual experiment, the lowest strength of cyanide solution required 
to extract an adequate percentage of the gold, and also to devise 
means of overcoming the problems which are inseparable from the 
treatment of different classes and grades of ore, with so active a 
compound as potassium cyanide. 

The daily output of ore from a mine is subject to continual 
change, both as regards physical condition and chemical composi- 
tion, hence the treatment requires, within certain limits, corre- 
sponding modifications, to obtain the maximum extraction at the 
minimum cost. To obtain these results, the metallurgist must be 
a trained analytical chemist, full of resource and originality. 

The testing and valuing of the ores before, during, and after 
treatment must be entrusted to a careful and trustworthy 
assay er. The testing and making up of the working solutions are 
very simple operations, which may be left to experienced and 
intelligent foremen who possess a knowledge of arithmetic extend- 
ing as far as decimals. 

With free-milling gold ores the actual working extraction will 
generally be as high as that obtained in the laboratory, but, in 
the case of ores containing copper or antimony, too much reliance 
must not be placed on the laboratory experiments. 

The author's experience is that high extraction may be obtained 
in the laboratory from ores totally unsuited for treatment by the 
cyanide process on a working scale. 

The conditions on the one hand are theoretical, on the other 
actual, and before adopting the cyanide treatment for a sulphide 
or other ore, working experiments should be made on parcels rang- 
ing from two to five tons, in order to ascertain the consumption of 




cyanide and actual extraction. If the working trials are successful 
the cyanide treatment may be adopted with confidence. 

On the other hand, in the case of an ore containing compara- 
tively coarse gold, the laboratory experiments — where the sample 
is hand-crushed — will give lower results than those obtained in 
practice in the cyanide works. The author made a number of 
experiments on an ore from Marlborough, N.Z. The average 
extraction in the laboratory was under 40 per cent., while the 
cyanide plant extracted over 60 per cent. At the battery the ore 
was dry-crushed through a 60-mesh screen, and investigation 
showed that a large portion of the gold was reduced fine enough 
to pass through the screen, and thus became amenable to cyanide 


1. Procure six bell-jars, about four and a half inches in diameter. 
When bell- jars are not procurable, lamp glasses or clear glass pint 
bottles, with the bottoms cut off, will answer the purpose quite 
well. In the neck of each jar fit a cork, perforated with one hole. 
Through the hole pass a short length of glass tube, on the end of 
which place a few inches of pliable black rubber tubing. On the end 
of the rubber tubing place a screw-clip, by means of which the rate 
of percolation of the cyanide solutions can be regulated to a nicety. 

2. Now invert the jars, and fix them in a wooden frame, so as 
to stand upright. In each jar place a thin layer of small rounded 
pebbles, about the size of French beans ; above the pebbles place 
an inch of coarse sand, and above this, half-an-inch of fine sand. 
Above the fine sand place a piece of loose scrim, the diameter of 
the jar. This completes the filter-bed. 

When a large number of cyanide experiments are being made, a 
box divided into three compartments, to hold the three grades of 
material for the filter-bed, should be kept well replenished and 
near at hand. 

3. Next procure a fair sample of the pulverized ore to be tested, 
weighing, say, six or eight pounds. Mix thoroughly, and carefully 
assay to accurately determine the original value. 

Check assays should always be made, and if there is a serious 
discrepancy between the assay and its check, amounting to over 
3 per cent, of the value, fresh assays should be made. The assays 
form the basis of the calculations and final results of the experi- 
ments, and hence the greatest accuracy should be aimed at. 

When the ore to be tested is from the battery, or mill, it should 
be placed in a jar in the condition that it comes from the mill, 


except, of course, when the tests are to determine the degree of 
fineness, which would give the best economic extraction. 

When the ore is hand-pulverized, a separate portion should be 
reduced to pass through, say, a 30-mesh, 40-mesh, and 60-mesh 
sieve respectively. Separate tests should be made of each grade, 
so as to determine to what extent the extraction is affected by the 
varying fineness of the ore. 

4. Introduce into each jar 10 or 12 ounces of the powdered and 
sampled ore, the value of which has been obtained by careful 
assay. Mark the jars 1, 2, 3, 4, 5, and 6. 

5. In the case of tailings or ores containing iron pyrites, or 
other base metallic sulphides, the samples in the jars should be 
washed once or twice with clean water to remove any soluble sul- 
phates. With very acid tailings a very dilute alkaline wash 
may be applied. (Test for acidity, see Chapter IV.) 

6. No general rule can be laid down as to the strength of the 
cyanide solutions to be used, as this will depend as much on the 
character of the sample as on its gold value ; but the strength of 
solution used on a working scale for any class of ore seldom 
exceeds 0*6 per cent. All the ores of silver, copper, arsenic, and 
antimony act on and consume cyanide, and when either of these 
is present a stronger series of solutions will have to be tried than 
in the case of clean ores. With the latter a useful series of 
solutions would contain : — 

005%, 0-1%, 0-15%, 0-2%, 0-25%, 0-3% of cyanide. 

With pyritic ores or tailings, or those containing copper, anti- 
mony, or arsenic compounds, the most instructive series would be : — 

0-15%, 0-20% 0-25%, 0-3%, 0-35%, 0-4%. 

In the event of all the tests being unsatisfactory, it would be 
advisable to try both stronger and weaker solutions than those 
already employed, according as the character of the ore or material 
may suggest. 

It is necessary with every new ore to make a number of laboratory 
experiments to ascertain the strength of cyanide solution required 
to extract an adequate proportion of the gold and silver contents. 

7. To each jar, already charged with the ore, add the same 
weight of solution as of ore. The excess of solution is required 
because a large portion immediately finds its way into the filter- 
bed. Record the numbers of the jars and the strength of cyanide 
used in each. Regulate the screw-clips so that the percolation 
will take at least thirty hours. A longer time may be tried if the 
first trials are not successful. With very dilute solutions, or when 
testing highly pyritic material, it may be necessary to continue 


the leaching for six days or even longer before satisfactory results 
are obtained. If the solution comes through too quickly, return 
it again and allow it this time to percolate more slowly. 

8. When the leaching is complete, wash with two washings of 
clean water, allowing the wash- water to drain a& rapidly as possible. 
The washing is complete when the wash-water no longer gives an 
alkaline reaction. 

9. Test the strength of the spent solution to ascertain the per- 
centage of cyanide used. The solution and washings are collected 
and measured together, then tested for cyanide. 

The consumption of cyanide can be calculated by determining 
the strength of the combined solution and washings, and making 
an allowance for the increase in bulk due to dilution. 

On a working scale the consumption of cyanide is generally 
much less than that shown by the laboratory experiments. 

Sometimes the cyanide and different washings are kept separate 
and evaporated down with the addition of litharge (in the manner 
described under The Assay of Cyanide Solutions), and the gold 
actually extracted by each, calculated separately. The results 
afford an instructive lesson on the value of successive washings. 

10. Remove the leached and washed residues from the jars, dry, 
mix thoroughly and again assay. As the residues will probably 
be excessively low-grade, it will be necessary to take 1000 or 1200 
grains for the assay determination. Then calculate the percent- 
age of extraction from each jar by difference, recording the results 
and assay values as follow : — 

Original Value. After Leaching. Percentage of 
ozs. dwts. grs. ozs. dwts. grs. Recovery. 

Gold, 2 4 12 4 12 89'9 

Silver, .10 4 3 6 83'9 

Value, .£9 £0 18 4 89*8 

The calculation is simply a matter of proportion. As an 
example take the recovery of gold. 

ozs. dwts. grs. 
Original gold, .2 4 12 per ton. 

After leaching, .0 4 12 „ 

Extracted, . .200 „ 

Then if two ounces were extracted from 2 ozs. 4 dwt. 12 grs., 
what would the extraction be from 100 1 



2 ozs. 4 dwts. 12 grs. : 2 ozs. : : 100 : 89*9 per cent. 

2 x 100 


= 89*9 per cent. 

An easy and expeditious method of calculating the percentages 
of extraction in the laboratory-test is to use the weights of the 
bullion, gold, and silver (in grains or grams) as the basis of com- 
putation instead of the same, extended as ounces, dwts., and 

Example : — 

Original After Extract ed. ^Z* 

Assay. Leaching. per cent. 

Bullion,. . '0020 '0002 -0018 900 






£9 2 £0 10 3 £8 11 9 



The calculation :- 

For Bullion. 

•0018 x 100 


For Silver. 

•0001 x 100 


= 90 

= 50 

For Gold. 
•0017 x 100 
•0018 ' 

For Value. 

8-587 x 100 


= 94*4. 

= 943. 

11. Compare the results obtained and adopt the strength which 
gives the highest extraction. 

Remarks. — With a series of experiments it will be found that 
the percentage of extraction, or rate of dissolution of the gold, 
reaches a maximum with a cyanide solution of a certain strength, 
and that above and below this strength the rate of extraction 
rapidly diminishes. The strength of cyanide solution which dis- 
solves the maximum percentage of gold will depend on the char- 
acter of the ore. The so-called selective action of cyanide is not 
so apparent in the laboratory experiments as it is in practice. 

On a working scale it soon becomes evident in the treatment of 
base sulphide ores that a strong solution of cyanide of potassium 
dissolves a large proportion of the base metals and a small propor- 
tion of the gold, while a weak solution dissolves a large proportion 
of the gold and a small proportion of the base metals. 



To Test the Strength of Cyanide Solutions.— This is an 

operation of great simplicity, and can be performed with accuracy 
and expedition by any intelligent foreman by a volumetric method 
of estimation. The standard solutions should alwavs be made up 
under the personal supervision of the chemist in charge of the works. 
Three different volumetric methods may be used for the deter- 
mination, namely : — 

1. By standard solution of silver nitrate. 

2. By standard solution of mercuric chloride. 

3. By standard solution of iodine. 

By Standard Silver Nitrate Solution. 

This is the method generally adopted in cyanide plants. It is 
a modification of Liebig's volumetric estimation of cyanogen. 

The reaction depends on the fact that when a solution of silver 
nitrate is added to a solution of potassium cyanide, the cyanogen 
unites with the silver, appearing as a white precipitate, which is 
immediately dissolved by any free KCy, which may still be pre- 
sent, forming a double cyanide of potassium and silver. 

This reaction is shown by the equations : — 

AgN0 8 + KCy - AgCy + KN0 8 ; 

AgCy + KCy=AgKCy 2 . 

A standard solution of silver nitrate can be made up from the 
molecular weights of the constituents as follows : — 

AgN0 8 saturates 2KCy. 

170 = 130 

17 = 13 

With grams, use a decinormal solution ; then if 17 grams of 
silver nitrate are dissolved in 1000 c.o. of water, 1 c.c. will be equal 
to -013 grm. of KCy. 


To Make Standard Silver Nitrate Solution. — Take 17 
grams of silver nitrate (triple crystallized if procurable), and 
dissolve in one litre (1000 c.c.) of distilled water. In large, 
works, where much testing is going on, it is advisable to dissolve 
34 grams in two litres ; then place in stoppered-bottle and mark. 

To Test Solutions of KCy : — 

1. Fill a burette with silver nitrate solution. 

2. Measure 13 c.c. of cyanide solution to be tested from 
another burette and transfer to a smaller beaker. To obtain . 
accurate results add a few drops of potassium iodide solution to 
the beaker and shake. 

3. Bun in standard AgN0 8 solution cautiously from the burette 
till the white precipitate formed just ceases to re-dissolve when the 
beaker is shaken ; that is, when a faint permanent opalescence 
appears the reaction is complete. 

4. Read off number of c.c. of standard solution used, and 
divide by 10. The result will represent the percentage of avail- 
able KCy. For example : — 

Suppose 13 c.c. of KCy sol. took 14*5 c.c. of AgN0 8l then — 

1^-1-46% of KCy. 

If a strong solution is being tested, in order to save AgN0 3 
measure off, say, 3 c.c. or 4 c.c. of cyanide solution instead of 13, 
and titrate with silver standard. Thus, if 4 c.c. required 6 c.c. 
of silver nitrate, 13 c.c. would require 19*5 ; and 19*5 divided by 
10 = 1-95% KCy. 

Even greater accuracy may be obtained in testing strong solu- 
tions, such as those in the dissolving tank, together with a saving 
of silver nitrate, by measuring off 13 c.c. of the strong solution 
and diluting with water to 130 c.c. Then measure off 13 c.c. of 
this diluted solution and titrate with silver nitrate as described 
above. Note the number of c.c. of standard solution required to 
complete the reaction, and this will represent the percentage of 
KCy in the strong solution, for since the 13 c.c. of dilute solution 
contained only a tenth of the original 13 c.c. of strong solution, 
there is hence no need to divide the quantity of silver nitrate by ten. 

To test the strength of very dilute cyanide solutions, measure 
off 130 c.c. of the solution, titrate with silver nitrate, and divide 
the number of c.c. of standard required by 100, and the result 
will give the percentage of available KCy, thus : — 

130 c.c. of cyanide solution required 5 c.c. of standard, then — 

5+ 100=0-05% KCy. 


To avoid calculation and reduce the liability to make mistakes 
in reading the burette, a standard solution of silver nitrate can be 
.made up by dissolving 13 '07 grams* of silver nitrate in 1000 c.c. 
of water. To test a cyanide solution with this, measure off 10 c.c. 
and titrate with silver nitrate. Note the number of c.c. of 
standard required to complete the reaction, divide by ten, and the 
result will be the percentage of available KCy, thus : — 

If 10 c.c. of cyanide solution required 5 c.c. of silver nitrate, 

5h- 10=0-5% KCy. 

Remarks. — Two burettes should always be used ; one to measure 
the cyanide solution, and one for the silver nitrate standard. 
The gram burette should be graduated to 1-1 0th gram. Erdmann 
floats should always be used, so as to obtain the exact reading. 

The addition of two or three drops of a 2 per cent, solution of 
potassium iodide to the solution to be tested renders the end 
reaction more defined, and reduces the danger of over-estimating 
the KCy due to the alkalinity of the solution. 

In an interesting paper on " The Titration and Use of Cyanide 
Solutions," t Mr. Walter H. Virgoe, Chief Chemist to the Mexican 
Gold and Silver Recovery Company, shows that where copper is 
present in the solution, the use of an iodide indicator fulfils the 
further purpose of indicating the point where the silver nitrate 
has titrated all the free cyanide, and is about to attack the 
cyanide in chemical combination with the copper, thereby pre- 
venting the over-estimation of the free cyanide. As an example, 
he says that a solution containing 0*3 per cent, of copper and 
titrating 0*52 per cent, of cyanide with silver nitrate alone, may 
titrate only 0*13 per cent, of KCy correctly if KI be used. 

Virgoe finds that in titrating pure solutions of potassium cyan- 
ide, the amount of indicator used makes no difference whatever, but 
when copper is present, even in small proportion, lie shows that 
with different amounts of indicator very dissimilar percentages of 
cyanide are obtained. For this reason he points out the advisa- 
bility of using a minimum addition of KI before titration of 
solutions containing copper. 

To Test Cyanide Solutions with Grain Standard Solution. 

1. When grain burettes are used, make up a solution of silver 
nitrate by dissolving 170 grains in 10,000 grains of pure water. 

2. Measure off, from a burette, 130 grains of cyanide solution 
to be tested. Add a few drops of KI. 

* Thus:— 17: 13 '08 : : 13:10. 

t Trans. Inst. Min. and Met., London, 1901-1902. 


3. Fill a burette with the silver nitrate standard. 

4. Run silver nitrate into cyanide solution until the white p.p. 
which at first forms just ceases to re-dissolve. Note the number 
of grains required. 

5. The number of grains of silver nitrate solution used to 
titrate, divided by 100, will give the percentage of available 

6. For example, if 56 grains of silver nitrate were used, then : — 


— — = 0*56 per cent, of KCy in the solution. 

By Standard Mercuric Chloride Solution. 

(1.) When a solution of mercuric chloride is added to a solution 
of potassium cyanide, a cyanide of mercury is formed, but is at 
once dissolved by any excess of KCy present. 

When all the free or available KCy has been used, a bluish- 
white opalescence of HgCy 2 appears if a slight excess of mercuric 
chloride is added. This permanent opalescence indicates the end 
of the reaction. 

(2.) To Make up Standard Solution. — Use the equation : — 

HgCl 2 + 2KCy = HgCy 2 + 2KC1. 

271 saturates 130. 

27*1 = 13 in a decinormal solution. 

From the above molecular weights, dissolve 27'1 grams of 
mercuric chloride in 1000 c.c. of distilled water; then 1 c.c. will 
equal -013 grm. of KCy. Place in a stoppered bottle and mark. 

(3.) The Actual Determination — 

(a.) Fill a burette with standard mercuric solution. 

(b.) From another burette measure off 13 c.c. of the cyanide 
solution to be tested, and to this add about 3 c.c. of 
dilute ammonia. 

(c.) Now run in standard mercuric solution very cautiously, 
with constant shaking, until a permanent bluish-white 
opalescence is produced. 

(d.) Note the number of c.c. of standard required to com- 
plete the reaction ; divide this number by 10, and the 
result will be the percentage of available KCy present ; 
thus, if — 
6*5 c.c. were required to complete titration, then — 

65 -r- 10 = 0*65% KCy. 


Remarks. — With pure substances this reaction is very delicate, 
but with cyanide solutions, containing much impurity, it is not so 
reliable as the silver nitrate method. Caustic alkalis do not 
interfere with the reaction. The author has made a number of 
simultaneous tests, with working cyanide solutions, by the silver 
nitrate and mercuric chloride methods, and the results obtained 
were practically the same throughout. 

By Standard Iodine Solution. 

(1.) This method depends on the fact that when a solution of 
iodine is added to one of potassium cyanide, the iodine loses its 
colour so long as any undecomposed cyanide remains. 

(2.) To Make up Standard Iodine Solution. — Use the re- 
action : — 

21 + KCy = KI + ICy. 

254 saturates 65. 

254 = 6*5 in a decinormal solution. 

Therefore, to make a standard solution, weigh out 25*4 grams 
of iodine, place in a beaker with 200 c.c. of water, and add suffi- 
cient potassium iodine to completely dissolve the iodine with 
frequent shaking. 

When the iodine is dissolved, make up to 1000 c.c. with pure 
water, and place in a stoppered bottle. Then : 

1 c.c. = '0065 grm. KCy. 

(3.) The Actual Determination — 

(a.) Fill a burette with the standard iodine. 

(b.) From another burrette measure off 6*5 c.c. of cyanide 

solution to be tested, and to this add carbonic acid 

(20 c.c. of ordinary soda water will do) to convert the 

caustic and mono-carbonate alkalis, contained in all 

commercial cyanide, into bi-carbonates. 
(c.) Now run in standard iodine, cautiously and slowly, 

until a slight but permanent yellow colour is produced. 
(d.) Read off the number of c.c. of standard employed, divide 

by 10, and the result will be the percentage of KCy 


Remarks. — This method does not give reliable results in the 
presence of sulphides, or when the cyanide solution is muddy or 

To Make up Cyanide Solutions. — There are two different 
methods of making up solutions in common use in cyanide plants. 


In some cases the requisite amount of solid cyanide salt is added 
to the sump-solution ; in others, the working strength is made up 
by adding strong solution from the dissolving tank to the sump 

The following exercises will render these methods clear : — 
1 lb. of pure KCy dissolved in 100 lbs. of water gives a 1 per cent, 
solution ; therefore, if you have a vat containing 100 cubic feet of 
water to make up to, say, 0*6 per cent., you would require 37*35 
lbs. of pure KCy.* 

Thus 100 x 62£ = 6225 lbs. of water, 

and if 100 lbs. of water require 0*6 lbs. KCy, 6225 lbs. would 
require — 

100 : 6225 : : 0'6 : x 

6225x0-6 o 7 .q*iv 

ioo - 37 35 lbs - 

Commercial cyanide is seldom pure ; you would, therefore, have 
to use a greater quantity to make up the required strength. 
Suppose the crude KCy contains 78 per cent, of KCy, then : — 

78 : 100 : : 34*35 : 7 
100 x 37-35 


= 48 lbs. crude KCy. 

The same form of calculation will do for making up any required 
quantity of cyanide solution. Suppose 4 ozs. of a 0*5 per cent, 
solution were required. 

Then if 100 ozs. of water require 0*5 ozs. of cyanide, how much 
would 4 ozs. require ? 

100 : 4 : : 0*5 : x 
^-^- 5 = 002 x 480 = 9-6 grains. 

If you have a 0*2 per cent, solution and you wish to make it up 
to, say, 0'5 per cent., subtract the 0*2 per cent, already in the 
solution from 0*5 per cent., leaving 0*3 per cent, required. Then 
proceed to make up as directed in the preceding paragraph. 


(1.) I have 4000 lbs. of sump solution containing 0*2 per cent. 

* See Constanta at end of Chapter. 


of available KCy, which I wish to make up to a 0*5 per cent, 
solution, how much additional KCy will be required ? 

100 : 4000 : : 0-3 : x 

4000 x -3 


= 12 lbs. pure KCy. 

If your crude KCy salt contains only 82 per cent, of KCy, 
then : — 

12xl00 aaU . 6 lbs crude KCy required 


(2.) How many lbs. of solid cyanide salt of 75 per cent, strength 
should be used to make up 10 tons of a 0*4 per cent, working 
solution? Ans. — 119*46 lbs. 

(3.) How many lbs. of solid cyanide salt of 82 per cent, strength 
should be used to make up 5 tons of a 0*45 per cent, working 
solution, using a sump solution containing 0*15 per cent, of KCy 
for making up? Ans. — 40*97 lbs. 

(4.) How many lbs. of a 14 per cent, stock cyanide solution 
should be used to make up 10 tons of a 0*4 per cent, working 
solution, using a 0*18 per cent, sump solution for making up? 

Solution. — This is easiest determined by " Alligation," the tea- 
mixer's rule of proportion, thus : — 

Strong solution 14 00 \S "40 working solution 
Working solution '40 /\ '18 sump solution 

Proportion of weak = 13*60 + *2 2 = proportion of strong. 
Neglecting decimals, then: — 

1360 + 22 = 1382. 

Here we have 1382 parts or lbs. of the required mixture, con- 
taining 22 of strong and 1360 of weak solution; therefore, if 22 
lbs. of the strong solution give 1382 lbs. of the required mixture, 
how many lbs. will be required for 10 tons of the mixture? 

1382: (10x2240) : : 22 : x 

22400 >x 22 = 356 , 58 

The answer is, therefore, 356*58 lbs. 

(5.) How many lbs. of a 22 per cent, stock solution should be 
used to make up 9 tons of 0*5 per cent, working solution, using a 
0*12 sump solution for making up ? Ans. — 351*27 lbs. 

(6.) How many lbs. of a twelve per cent, stock solution should be 
used to make up 10 tons of a 0*6 per cent, working solution, using 


0*15 per cent, sump solution for making up? Before drawing 
from the stock solution first utilize 4 tons of a 0*8 solution already 
in the solution vat. Ans. — 359*74 lbs. 

Solution of No. 6. — First find out how much of the working 
solution can be made up from the 4 tons of 0*8 per cent, cyanide 
solution, thus: — 

Strong solution '80 \ / '60 working solution 

>< ; ' 

Working solution *60 /\ '15 sump solution 

Proportion of weak = '20 + '45 = proportion of strong. 
Neglecting decimals, then : — 

20 + 45 = 65 of required mixture. 

Then, if 45 of the strong (0*8 per cent.) give 65 of the required 
mixture, the 4 tons already in the solution tank will give 5*77 
tons, thus : — 

45 : 4 : : 65 : x 



= 577. 

And 10 — 5*77 = 4*23 tons to be made up from the 12 per cent, 
stock solution, thus : — 

Strong solution 12*00 \/ '60 working solution . 
Working solution *60 /\ *1 5 sump solution 

Proportion of weak = 1 1 '40 + -45 = proportion of strong. 
Neglecting decimals : — 

1 140 + 45 = 1 185 of mixture. 
Now, if 45 lbs. of the stock solution (12 per cent.) give 1185 
lbs. of the required mixture, how much will 4*23 tons require ? 

1185: (4-23x2240) : : 45 : x 

4*23 x 2240 x 45 = 359 . 741bg 
, 1185 


(7.) How many tons of a 0*45 per cent, working solution would 
6 tons of a 0'8 per cent, solution of cyanide make, using water for 
dilution? Ans. — 10*66 tons. 

Solution : — 

•45 : *80 : : 6 : x 

8 _0|_6 = 10 .66. 


(8.) How many tons of a 0*4 per cent, working solution of 
cyanide would 8 tons of a 0*6 per cent, solution make, using a 0*12 
per cent, sump solution for dilution? Ans. — 13*71 tons. 

Solution : — 

Strong solution "60 \ / "40 working solution 

><- : 

Working solution "40 /\ "12 sump solution 

Proportion of weak = *20 + *28 = proportion of strong 

20 + 28 = 48 of mixture. 

Now, if 28 of the strong solution give 48 of the required mix- 
ture, 8 tons will give 13*71 tons, thus: — 

28 : 8 : : 48 : x 

^i? = 13*71 tons. 


(9.) How many tons of a 0*6 per cent, cyanide solution would 8 
tons of a 0*7 per cent, solution make, using a 0*2 per cent, sump 
solution for dilution ? Ans. — 10 tons. 

To Test the Strength of Crude KCy. — Commercial KCy is 
formed when any nitrogenous organic bodies, such as hoofs, clippings 
of hides, wool, and blood are fused with potassium carbonate. This 
product is very impure, and is lixiviated in a vessel containing 
finely-divided metallic iron, yielding the yellow prussiate of potas- 
sium (K 4 Fe 6 Cy 6 ), which is the starting-point of all cyanogen 

Crude cyanide of potassium is formed by the action of heat on 
the yellow prussiate, thus : — 

K 4 FeC 6 £T 6 = 4KCN + FeC 2 + N 2 . 

The chief impurities in commercial cyanide are black carbide of 
iron, alkaline carbonates, and sometimes alkaline chlorides and 
sulphides in small quantities. 

To accurately test the strength of the solid cyanide salt, for the 
free or available KCy which it contains, proceed as follows : — 
(1.) Break a cake of KCy in two, and select a piece, say, a pound 

in weight, showing the whole thickness of the cake in 

(2.) Reduce this pound to a coarse powder, sample well, and 

f urthur pulverize to a moderately fine powder. 
(3.) Weigh out 1 gram of powdered and sampled KCy. 
(4.) Dissolve in pure water and make up to 100 c.c. 
(5.) Measure off 13 c.c. of this solution and titrate with silver 

nitrate standard solution from a burette as previously 

described. Note number of c.c. of standard required to 


form a permanent p.p. ; divide by 10 and this will give 
the amount of KCy in 1 gram of the crude salt. 
For example : Suppose 13 c.c. of KCy solution required 7*5 c.c. 
of standard, then : — 


— — = *75 KCy in 1 gram, 

which is equal to 75 per cent, of KCy in the crude salt. 


The estimation of the gold contents of cyanide solutions in large 
Yi'orks where a great many determinations are made daily must 
be effected by a method both expeditious and accurate. Several 
alternative methods are given below, all of which give reliable 
results with ordinary care. Method I. is in very common use. 
It has the advantage over methods II. and III. that it can be 
used for the determination of both gold and silver contents, which 
is necessary in the case of argentiferous gold ores. Methods II. 
and III. are silver nitrate processes devised by Andrew F. Crosse, 
the well known South African metallurgical chemist. Method 
IV. is a copper sulphate process used by Walter H. Virgoe for 
some years for the rapid assaying of cyanide solutions. It is 
recommended by Crosse. Like method I. it can be used for 
valuation of gold and silver contents. 

Method I. — 

(1.) Measure half a pint of solution and evaporate slowly to a 
small bulk in a round iron drying-dish, over a Bunsen 
flame, or on the furnace lid. As the evaporation proceeds, 
rub the sides down so as to collect the whole of the dis- 
solved salts at the bottom. 

(2.) To the solution add 600 grains of litharge. Mix well; 
evaporate cautiously to dryness. 

(3.) Then transfer to a clay-crucible and mix with 200 grains of 
glass-powder, 100 grains of soda, and 48 grains of argol 
(bi-tartrate of potash). Cover with a little borax, and 
fuse. When fused, pour and allow to cool. 

(4.) Cupel the lead-button and weigh the resulting bead of 
bullion. If the ore contains silver, part so as to determine 
the weight of gold and of silver ; then refer to the table at 
the end of the chapter to ascertain the quantity of each 
per ton of solution. 

When the resulting gold is weighed with gram weights, refer to 


the Gram Table ; and, when in grains to the Grain Table. (See 
end of chapter.) 

Remarks. — When a large number of determinations have to be 
made, ordinary enamelled plates and mugs form efficient evaporat- 
ing dishes ; in this case also the litharge can be stirred into the 
solution before the evaporation begins. 

At many cyanide works the cyanide solutions are assayed by 
evaporating a measured portion of the solution in a boat of sheet- 
lead and then scorifying the residue, and cupelling. When the 
solutions are charged with base metallic cyanides the results are 
not generally so reliable as when the solution is evaporated with 
litharge and afterwards fused in a clay crucible. 

Method II. (Crosse). — 

(1.) Measure half a pint of cyanide solution, and add silver 
nitrate solution until a precipitate ceases to form. The silver salt 
should be added a little at the time, and the solution well shaken 
after each addition. All the gold in the solution is precipitated 
as argentic-auric-cyanide. 

(2.) Allow the precipitate to settle ; decant off the clear solu- 
tion; filter and dry the precipitate and mix with 200 grains 
litharge, 100 grains glass-powder, 100 grains soda, and 48 grains 
of argol. Fuse, pour, and cupel the lead-button. 

(3.) Extract the bead of bullion from the cupel, flatten, and 
part without weighing. 

(4.) Weigh the resulting gold and calculate the results. 

Method III. (Crosse).*— 

(1.) Pour 500 c.c. of cyanide solution into an evaporating dish. 
Put it in a stink cupboard with a good draught. 

(2.) Add nitric acid till the solution shows an acid reaction. 

(3.) Boil 15 minutes. 

(4.) Then add £ gram of silver dissolved in silver nitrate. 

(5.) Filter ; fuse the filter-paper and contained precipitate as 
usual with litharge, and flux; then cupel and weigh resulting 

Method IV. (Virgoe).— 

(1.) To a litre of solution add excess of weak sulphate of copper 

(2.) Acidify with hydrochloric, nitric, or sulphuric acid. 

(3.) Filter. The precipitate, which is white and flocculent, 
contains all the gold and silver. The bluish or greenish colour of 
the filtrate indicates that excess of copper sulphate has been added. 

* The Journal of the Chemical and Metallurgical Society of S.A., May, 


(4.) Wash precipitate, dry and place in a scorifier with metallic 

Crosse prefers to fuse the precipitate in a crucible with litharge 
and to cupel resulting lead button. 

Notes on Virgoes Method. — In a discussion on his method 
before the Chemical and Metallurgical Society of South Africa, 
Mr Virgoe supplied the following useful particulars : — 

"In dealing with ordinary solutions, I add copper sulphate 
carefully until no further precipitate forms, stirring the solution 
in the meanwhile, and taking care to add only a very slight 
excess of copper sulphate, attested by the faint tinge of blue in 
the filtrate. Allow the precipitate to stand for a minute or two, 
H 2 S0 4 is then added till the solution is acid, no more, no less. 
An addition of sodium sulphate produces in the filtrate a slight 
additional* prepicitate which is found to contain a very small 
amount of gold. I find the results are very accurate. For 
instance, 1 litre of solution was found to contain by careful 
duplicate evaporation tests, 8*27 milligrams; the copper sulphate 
method run in duplicate yielded 8*22 and 8*23 milligrams, just 
a little low, but the test is severe and amply demonstrates to my 
mind the correctness of the method if properly run. 

" As far as I have observed there seems to be a tendency, if 
strong solutions of copper sulphate are added to strong solutions 
of cyanide, towards the formation of cupric salts, henoe I use 
weak solutions of copper sulphate, and should expect bad results 
if the precipitation of gold were attempted by my method from a 
very strong solution of cyanide with a strong solution of copper 
sulphate.' 1 


Working solutions of potassium cyanide soon become con- 
taminated with ferro-, ferri-, and sulpho-cyanides and cyanates 
of the alkaline earths and base metals soluble in KCy, some of 
which have a beneficial and others an injurious effect. In order, 
therefore, to be able to modify the treatment to obtain a higher 
rate of extraction, or to reduce the consumption of cyanide, or effect 
a better precipitation, it is necessary to know something of the 
constitution of the working solutions which generally contain, 
even in the treatment of the purest ore, an assemblage of com- 
plicated, compounds requiring much technological skill for their 

The analysis of cyanide solutions is an object to which many 
accomplished chemists, mostly in South Africa, have devoted 



much attention. Many useful schemes have been formulated, 
the object in most cases having been to devise reliable methods, 
devoid of too much refinement, yet capable of every-day applica- 
tion in cyanide works. 

The " Shaking Test " for Consumption of Cyanide.— 

This method is much used in the laboratories of the Cassel 
Cyanide Company, and affords a rapid and fairly approximate 
estimate of the consumption of cyanide with different classes of 
ore. It is useful for comparative purposes, and as a preliminary 
means of determining the most suitable strengths of cyanide 
solutions for laboratory experiments. 

(1.) Take 200 grams of the ore and place in a stoppered bottle 
with, for example, 100 c.c. of a 0'5 per cent, cyanide solution, and 
shake for twenty minutes. 

(2.) Allow contents of bottle to settle ; draw off a portion of the 
clear solution with a pipette and test for KCy. If it contains 0*2 
per cent, of KCy, then 0*3 per cent, has been consumed or decom- 

(3.) When much cyanide is used up, test the ore for acidity by 
Feldtmann's method given below. 

To Test Ores and Tailings for Acidity. — 

(1.) Weigh out 224 grams of ore and shake up with 250 c.c. of 
water in a tall glass-jar or cylinder. 

(2.) Fill a burette with a standard solution of soda, and titrate 
the ore-solution in the jar until the reaction is neutral to test 
(litmus) paper. 

(3.) Every c.c. of the soda solution used will represent 0*1 lb. 
of caustic soda to be added to every ton of ore (or tailings) in a 
wash before the cyanide treatment. 

To Make Standard Soda Solution. — Dissolve 10 grams (or 
154*3 grains) of caustic soda in 1000 c.c. of pure water, and place 
in a secure bottle. 

Remarks. — During the titration, the litmus paper should be 
dipped in clean, pure water from time to time to remove the 
adhering particles of ore so that the reaction may be clearly seen. 

Tests for Alkaline Sulphides in Cyanide. — Alkaline sul- 
phides act injuriously in cyanide solutions during leaching, and it 
is important to detect their presence. They are all soluble in 

First Test : — To the clear cyanide solution add a little acid. If 
an alkaline sulphide be present, sulphur will be liberated, impart- 
ing a cloudy appearance to the solution. 

Second Test : — In the clear solution place a clean, bright silver 
coin. It will become black and tarnished if a sulphide be present. 
This and the preceding test will not detect minute quantities. 


Third Test : — The most delicate test is by means of the nitro- 
prussides. These are formed by adding a little nitric acid to a 
solution of ferro- or ferri-cyanide of potassium. 

Add a few drops of a solution of nitro-prusside to the cyanide 
solution. If an alkaline sulphide is present, even in minute 
quantity, the solution will assume a brilliant purple colour. 

Fourth Test : — When a dilute solution of a soluble lead salt, 
such as the acetate, is added to a solution containing an alkaline 
sulphide, a blackish brown precipitate of lead sulphide soon forms. 

When the alkaline sulphide exists in the presence of free 
cyanide, a white precipitate of lead cyanide and carbonate will 
immediately form on the addition of lead acetate, thus tending to 
render the lead sulphide precipitate yellowish or nut brown. 

(Feldtmann and Bbttel.) 

From 5 to 10 or more grams of the commercial cyanide are 
dissolved in water, and the insoluble matter, if any, filtered off. 

The solution is agitated with a small quantity of precipitated 
lead carbonate — of course slightly in excess, and filtered. The 
precipitate, consisting of lead carbonate and sulphide, is trans- 
ferred to a flask, and covered with a few c.c.'s of a solution of 
potassic or sodic cyanide free from sulphides, sulphocyanides, or 
ferrocyanides. This solution may be prepared from pure potassic 
or sodic hydrate and a solution of pure distilled hydrocyanic acid. 

To the mixture in the beaker add hydrogen peroxide in slight 
excess — i.e., three or four times as much as is needed to whiten 
the precipitate. (The hydrogen peroxide for this purpose should 
be purified by agitation with ether and evaporation of the ether 
in a water-bath.) 

A small quantity — say \ gram — of manganese peroxide is then 

added, and the mixture agitated for about two minutes, after 

which the solution is filtered off, acidified with sulphuric acid, 

and titrated with — — potassic permanganate. 

1 c.c. of potassic permanganate equals 0*000053 grm. sulphur 

or 0*000182 grm. potassic sulphide. 

The potassic permanganate may be standardized by means of pure 

potassic sulphocyanide. 

1 c.c. = -0001618 grm. KCyS. 
* Paper read before the Chemical and Met Society of S. A. 


Estimation of Hydrocyanic Acid.— -To 50 c.c. of the 

solution add a solution of bicarbonate of potash or soda, free from 
carbonate, or excess of carbonic acid. Titrate as for KCy. 
Deduct from this the amount of KCy found. The difference 
equals the amount of HCy present. 

1 c.c. AgN0 3 = 0-0414 HCy. 
Estimation of double Cyanides.— Add excess of pure 

caustic soda to 50 c.c. of solution, and a few drops of KI solution, 
then titrate with AgN0 3 . Deduct KCy and HCy, as found 
above ; the difference is K 2 Zn(Cy) 4 as KCy*, less 7 '9 per cent. 

The KCy as found here is calculated to K 2 Zn(Cy 4 ), as 
under : — 

KCy x 0-9493 = K 2 Zn(Cy) 4 . Add to this 7 '9 per cent, of total, 
or for every 92 -1 parts add 7*9 parts. 

Estimation of Ferro- and Sulpho-Cyanides.— When 

organic matter is present, shake with powdered quicklime and 

A burette is filled with the cyanide solution for analysis, and 
run into 10 or 20 c.c. 1 / 100 Normal K 2 Mn 2 8 strongly acidified 
with H 2 S0 4 until the colour is just discharged. 

A solution of ferric sulphate or chloride is acidified with H 2 S0 4 
and 50 c.c. of the cyanide solution poured in. After shaking for 
about half a minute the Prussian blue is separated from the 
liquid by filtration, and the precipitate washed. The filtrate is 
next titrated with 1 / 100 normal K 2 Mn 2 8 . 

Let x be c.c. permanganate required to oxidize ferro-cyanide, 
then x = y - z. 

(x) 1 c.c. Vioo normal KjM^Og = 0'003684 grm. K 4 Fe(Cy) 6 . 
(z) 1 c.c. Vioo normal K^Mo^O^ 0*0001618 grm. KCyS. 

(by Leonard M. Green, A.R.S.M.) 

The tests used depend essentially on alkalimetric determinations. 
In Mr. Green's scheme the following constituents are estimated: — 

(1.) The total cyanide =T. 

(2.) The protective alkali =p. 

(3.) The alkaline or alkaline-earth hydrates = h. 

(4.) Alkaline mono-carbonates =N. 

(5.) The ferrocyanides =S. 

(6.) The zinc = Z. 

* Extracts from Paper read before the Inst. Min. and Met., London. 


The methods depend on the facts that : — 

(I.) Potassium ferrocyanide is neutral to phenolphthalein. 

(2.) That I c.c. of decinormal potassium ferrocyanide pre- 
cipitates 0*75 c.c. decinormal zinc from a neutral dilute 
solution of a zinc salt. 

(3.) That when a dilute neutral solution of a zinc salt is 
precipitated by adding an excess of sodium carbonate 
solution, the excess of alkali being afterwards carefully 
neutralized to phenolphthalein, by the addition of 
decinormal acid, a precipitate of basic zinc carbonate of 
almost constant composition is obtained. The precipi- 
tate obtained in this way is the normal basic carbonate. 

(4.) That zinc hydrate or carbonate, when treated with an 
excess of potassium ferrocyanide, forms zinc ferro- 
cyanide and potassium hydrate or carbonate, the 
alkalinity produced being proportional to the precipitate 
acted on. This reaction does not immediately proceed 
to the end, as at first only a portion of the alkalinity is 
formed ; but if this be neutralized with acid a further 
amount of alkali is formed, and so on to the finish, the 
reaction taking a little time for completion. 

The Actual Analysis. — (1.) The first estimation is that of 
the total cyanide. This is performed in the usual way by adding 
to the solution to be tested an excess of caustic soda and a little 
potassium iodide, and titrating with silver nitrate till a distinct, 
permanent yellowish cloudiness appears. It must be noted that 
the end-point is not reached until there is a distinct yellowish 

Where much zinc and ferrocyanides are present, a faint, white 
cloudiness, probably due to the precipitation of zinc-ferrocyanide, 
is sometimes produced before the true " end-point." This must 
be disregarded, the true " end point " occurring when the yellowish 
cloudiness, due to silver iodide, is permanently formed. This is the 
only definite end-point in titrating such a solution, and a large 
excess of sodium hydrate does not appreciably alter it. 

(2.) In the second test the alkaline and alkaline-earth hydrates 
plus half the mono-carbonates, viz., the "protective alkali," is 
determined. This test is a simple alteration of Clennel's. 

Excess of potassium ferrocyanide is added to the solution, and 
then twice the amount of silver nitrate necessary to indicate the 
total cyanide, viz., sufficient to precipitate the whole of the 
cyanide. A slight excess does not matter, as it merely precipi- 
tates some chloride, or, if no chlorides are present, some sulpho- 
cyanide or ferrocyanide. The zinc all occurs in the precipitate as 


ferrocyanide, and the alkaline hydrates and carbonates are left in 

Phenolphthalein is then added, and the solution titrated with 
decinormal nitric acid till colourless, or till it acquires the faint 
greenish yellow tinge produced by the excess of ferrocyanide. 
The result indicates the "protective alkali." 

Usually an identical result is obtained by only adding a little 
more than the amount of silver nitrate necessary to indicate the 
total cyanide, but this leaves potassium silver-cyanide in solution, 
and towards the close of the acid titration there is a slight chance 
of its being acted on by the acid forming AgCy, and setting free 
HCy, which would slightly increase the acid required, and obscure 
the end-point. 

(3.) In the third test the alkaline hydrates are estimated. An 
excess of barium chloride is first added to the solution (sufficient 
to precipitate the sulphates and carbonates), and then the proced- 
ure in the last test is repeated. The result obtained indicates 
only the alkaline hydrate. 

In the event of no hydrates being present, bicarbonates probably 
exist in solution. They may be estimated by adding a known 
amount of standard sodium hydrate, and repeating the test, allow- 
ing, in calculating the results, for the amount of hydrate added. 

(4.) The next estimation is that of the total cyanide + chlorides 
+ sulphocyanides + ferrocyanides + any other salt precipitated 
by silver nitrate before the precipitation of chromate. This test, 
though not important in itself, is necessary for the subsequent 
determination of the zinc. 

The amount of acid used in the second test is first added to the 
solution (viz., sufficient to neutralize the protective alkali). Then 
one drop of a strong solution of potassium chromate is added, and 
the solution is titrated with silver nitrate till there is a faint per- 
manent reddish colouration. 

Towards the end of the titration, the reddish colouration appears, 
and only goes again slowly on shaking and standing a few seconds. 
This is due to the fact that the action of the silver nitrate on the 
precipitated zinc ferrocyanide is slow, and consequently some silver 
chromate is temporarily formed; this chromate, however, when 
well shaken up with the zinc ferrocyanide, is decomposed, and 
silver ferrocyanide is formed. The end-point therefore only occurs 
when the reddish colouration is decided and permanent. 

(5.) The fifth estimation is that of the zinc - the ferrocyanide. 
Sufl&cient sodium carbonate is added to supply enough carbonic 
acid for the basic zinc precipitate in case there should not be 
enough already in the solution. Then twice the amount of silver 
nitrate necessary to indicate the total cyanide is added and the 


whole is well shaken. All the cyanide is now precipitated as 
silver cyanide, and any slight excess of silver nitrate will merely 
precipitate some chloride. 

The zinc will have been partially precipitated as ferrocyanide 
by the ferrocyanide originally present in solution, but the remain- 
der will have been precipitated as a basic oarbonate, and there 
will be an excess of alkali in the solution. This alkalinity is 
neutralized by the addition of decinormal acid. During this 
neutralization the basic zinc carbonate gradually acquires its 
normal composition 3Zn (HO) 2 .2ZnC0 3 , and the neutralization 
must not be hurried. 

The colour of the clear solution above the precipitate must not 
show a trace of pink, even on standing a minute or two, aud the 
contents of the flask should be frequently shaken. In fact, when 
neutralization is apparently complete, it is best to add another c.c. 
of acid, shake up well and then add from a burette, a drop or so 

at a time, a solution of sodium carbonate (roughly — - is a con- 

. venient strength), till the clear solution again just shows a pink 

All the zinc which is not precipitated as ferrocyanide is now in 
the precipitate as basic carbonate, and the solution is neutral to 
phenolphthalein. If now an excess of potassium ferrocyanide be 
added, the basic zinc carbonate reacts with the ferrocyanide, 
forming zinc ferrocyanide and alkalinity proportional to the 
amount of basic carbonate present. 

This alkalinity is then titrated with decinormal acid, the result 
representing the zinc less what has been precipitated by the 
ferrocyanide originally present. 

Towards the end of the titration the colour is discharged and 
returns again slightly, so that a little time is necessary for the 

(6.) The last determination is of the zinc. Sodium carbonate is 
added as in the previous test ; then silver nitrate solution is run 
in, in the same amount as was used in test No. 4, and the flask 
well shaken. The cyanides, chlorides, sulphocyanides and ferro- 
cyanides will then have been precipitated as silver salts, and the 
whole of the zinc will have been precipitated as a basic carbonate, 
an excess of alkali remaining in the solution. This is naturalized 
precisely as in the last test, an excess of ferrocyanide added, and 
the alkalinity produced titrated with decinormal acid. 

This result consequently represents alkalinity proportional to 
the whole of the zinc. It includes metals acting similarly to 
zinc, such as cadmium, and copper if present in small quantity. 


Summary. — The following is a short summary of the tests 
described, with the quantities of solution taken, etc. : — 

(1.) Take 50 c.c. of the solution. Add excess of sodium hy- 
drate and a little potassium iodide solution. Titrate with silver 
nitrate till there is a distinct permanent yellowish cloudiness. 

Kesult = T c.c. 

(2.) Take 50 c.c. of the solution. Add excess of potassium ferro- 
cyanide solution. Kun in 2T c.c. of silver nitrate. Add phenolph- 

thalein and titrate with — — nitric acid till colourless. 

Result = p c.c. 

(3.) Take 50 c.c. of solution. Add excess of BaCl 2 solution, and 

then excess of potassium ferrocyanide. Run in 2T c.c. of silver 

nitrate, add phenolphthalein and titrate with — nitric acid till 


Result = h c.c. 

(4.) Take 50 c.c. of the solution. Add 2T c.c. of silver nitrate, 

then p. c.c. of — - nitric acid. Add one drop of a strong solution 

of potassium chromate, and continue titrating with silver nitrate 
till there is a faint permanent reddish colouration. 

Result (Total AgNo 8 added) = Nc.c. 

(5.) Take 50 c.c. of solution. Add about 10 c.c. of sodium car- 
bonate solution (roughly decinormal). Run in 2T c.c. (see Test 1) 
of silver nitrate, and shake well. Add phenolphthalein, and 

neutralize with — - nitric acid till the clear solution is colourless. 

Shake well at intervals during the neutralization. Add about 

1 c.c. more -— nitric acid, shake up, and then add a solution of 

sodium carbonate (roughly -- - is a convenient strength), drop by 

drop, till the clear solution is just faintly pink. 

Add excess of potassium ferrocyanide. The solution becomes 

strongly alkaline to phenolphthalein. 

Titrate with — - nitric acid till colourless. 

Result = S c.c. 


(6.) Take 50 c.c. of solution. Add about 10 c.c. of sodium car- 
bonate solution. Run in N c.c. of silver nitrate (see Test 3), and 
shake well. 

Add phenolphthalein, and neutralize as in the previous test. 

Add excess of potassium ferrocyanide. The solution becomes 
strongly alkaline to phenolphthalein. 

Titrate with — - nitric acid till colourless. 

Result =Z c.c. 

Notes. — If the standard solutions used are — nitric acid, and a 

solution of silver nitrate containing 13*05 gram per litre, then 
the following are the factors for the different results : — 

(1.) T x 0-02% = Total Cyanides estimated as KCy. 

(2.) p x 00112% = Protective Alkali estimated as KHO. 

(3.) h x 0*0112% = Alkaline hydrates estimated as KHO. 

(4.) p - h x 0*0276% = Alkali carbonates. 

(5.) Z-Sx 0*0351% = Potassium ferrocyanide. 

(6.) Zx 00081% = Zinc. 

The presence of sulpho-cyanides or ferrocyanides in small 
quantity does not interfere with these tests. The latter in large 
quantity does, as precipitates of zinc and silver ferrocyanides 
rapidly destroy the indicator phenolphthalein although potassium 
ferrocyanide itself does not appear to do so. 

A large excess of sodic carbonate has hardly any effect on the 
final result of the titration for zinc, but it tends to decrease the 
sharpness of the end points, and increases the time taken for the 


(By Andrew F. Ceosse.) 

Prepare a solution containing : — 

20 grams Iodide of Potassium, 
2 .„ Nitrate of Potassium, 

dissolved in distilled water and made up to 100 c.c. 

The sample of cyanide solution can be taken in an ordinary 
Winchester quart bottle filled perfectly full. 

Syphon off a sample into a separating funnel containing from 
250 to 300 c.c, then add 1 c.c. of the solution of iodide and 

* Jour. Chem. and Met. Soc, S.A., May 1902. 


nitrate of potassium, and 3 c.c. of dilute sulphuric acid (1-1). 
After shaking up and allowing to stand for 15 minutes, titrate in 
an atmosphere of coal gas with a hyposulphite of sodium solution 
containing 7*75 grams per litre, 1 c.c. of which is equal to 0*25 
m'grams of oxygen for the iodine liberated. 

The peculiarity of the reaction is that no iodine is apparently 
liberated, and the cyanide solution remains colourless. But iodine 
has been liberated all the same to form a very unstable compound, 
iodide of cyanogen, and the amount is in exact proportion to the 
free oxygen plus the quantity to be subsequently allowed for, and 
can be estimated by the hyposulphite solution, using a starch 
solution as an indicator. Of course it is necessary to estimate the 
correction required for the iodine liberated by the reagents, and 
also in case any nitrates were present in the cyanide solution. 

This correction is easily made in the following manner : — 

Take about 400 c.c. of the solution under examination, add to 
it 0*3 grams pure ferrous sulphate and the same weight of 
caustic lime, shake up well and filter into a flask through which 
coal gas is passing, the precipitated ferrous hydrate will have ab- 
sorbed all the free oxygen in the cyanide solution. 

A good many experiments were made to test the above method 
and the following will be a good example : — 

A Winchester quart bottle was partly filled with a sample of an 
ordinary working solution, the bottle was well shaken for some 
time in order to saturate the solution with oxygen. 

The determination was made as described, using a pipette con- 
taining 292 c.c. of solution, 13*4 c.c. of hyposulphite of soda was 
required to unite with the iodine liberated. 

Then 400 c.c. of the cyanide solution was freed from oxygen 
and the amount of hyposulphite required was 6*0 c.c. 

So that 13*4 — 6*0 leaves 7*4 c.c. of hyposulphite required for the 
iodine liberated by the free oxygen in 288 c.c. of solution or 1*85 
m'grams oxygen in 288 c.c, or 6*4 m'grams oxygen per litre. 

In connection with the oxygen determination Mr. Crosse has 
made some interesting experiments on the amount of oxygen ab- 
sorbed by sands and spitzluten concentrates. 

He says : " I take, say, from 200 to 500 grams of the material 
under examination and shake it up for some hours in a large 
bottle full of a cyanide solution of known quantity and saturated 
with oxygen and then determine the amount absorbed or lost per 
litre, and so calculate out the amount of oxygen required by each 
kilogramme or ton of sands. I would, however, remark that 
though oxygen is being absorbed by certain substances during 
treatment, such as ferrous sulphides, etc., gold is also being dis- 
solved, but more slowly and to a lesser degree than if no oxygen 


absorbing matters were present. If we have ferrous hydrate 
present, which would almost immediately take up the oxygen, 
then no gold would be dissolved, but certain substances act slowly, 
allowing some gold to be dissolved at the same time that oxygen 
is being absorbed. If this were not the case it would often happen 
that no gold would be dissolved till several solutions had passed 
through the sands." 



(By Andrew F. Crosse.) 

The following is an accurate method for determining copper, 
iron, zinc and nickle in solutions : — 

Take from 500 to 1000 c.c. of solution, acidify with a slight 
excess of sulphuric acid, add five or six grams of pure acid 
sulphate of potassium, and evaporate to dryness in a platinum 
dish and then heat to dull redness in order to melt the mass. 
The metals are obtained as sulphates and can be separated and 
estimated in the usual way. 


In the solutions on the Rand, where we are not usually troubled 
with anything more than a trace of copper, the chief metallic 
ingredient is necessarily zinc and the estimation of this metal 
would often be useful. 

Take 300 c.c. of solution, add about a gram of cyanide of 
potassium and the same quantity of pure caustic potash or soda, 
heat nearly to boiling point and then add a slight excess of 
sulphide of sodium in solution. The zinc will be quickly pre- 
cipitated as a sulphide, and should be collected on a filter paper 
and washed with hot water. Then place the filter paper in a wide 
mouthed bottle of known capacity between 250 and 300 c.c. 

This bottle must be provided with a well fitting india-rubber 
bung through which a moderately wide tube is inserted about 8 
to 10 inches long. 

Then fill up the bottle with a weak solution of pure ferric 
sulphate containing 5 to 7 per cent, of sulphuric acid, place the 
bottle or flask in a bowl of cold water and raise the temperature 
to boiling point. The reason for the glass tube will be apparent 
as it allows for the expansion of the liquid. 

*Jour. Chem. and Met. Soc., S.A., May 1902. 


The zinc sulphide will have decomposed and reduced a pro- 
portionate amount of ferric sulphate to ferrous sulphate. When 
nearly cold, filter off the solution through a dry filter paper, and 
take half the quantity contained in the bottle and titrate with 
decinormal permanganate of potassium. 

1 c.c. = "00325 grams of zinc. Sulphide of zinc is, however, 
slightly soluble in weak cyanide solutions, but having made 
various experiments, it was found that one milligram of zinc 
per 100 c.c. of solution taken, would be the right amount to add 
as a correction to the results obtained. 

Various solutions were prepared containing known quantities of 
the double cyanide of zinc and potassium with ferro-cyanide and 
sulpho-cyanide of potassium, and the results obtained were very 


(By Andrew F. Crosse.) 

Pure cyanide of potassium becomes rapidly decomposed if 
brought into contact with air, owing to the presence of carbonic 
acid. Most South African tailings and slimes contain a certain 
small percentage of free sulphuric acid, or basic ferric salts, which 
decompose cyanide of potassium. We must protect or prevent 
this decomposition as far as possible. Cyanide of potassium is an 
expensive material; we should therefore endeavour to prevent 
its unnecessary decomposition, by the use of some inexpensive 
alkali — such as lime — and it becomes necessary to have an 
accurate method for determining the amount of caustic lime or 
other protective alkali in solution. 

Protective alkali is the alkali present which will unite with any 
acid before decomposition of the cyanide begins. 

It is very important in many cases to be able to determine the 
percentage of this alkali exactly. 

In my experiments lime (calcium oxide) is taken as the protec- 
tive alkali, but the results can be calculated out for caustic potash 
or caustic soda, if so desired. 

If an exact amount of sulphuric acid be added to a solution of 
pure cyanide of potassium containing the latter in excess, we have 
the following reaction : — 

2KCN + H 2 S0 4 = K 2 S0 4 + 2HCN or 98 parts H 2 S0 4 liberate 
54 parts of hydrocyanic acid. 

Now, if instead of H 2 S0 4 we take pure bi-sulphate of potassium, 
the following reaction takes place : — 

* Trans. Chem. and Met. Soc, S.A., 1899. 


KHS0 4 + KCN = K 2 S0 4 + HCN or 136 parts KHS0 4 liberate 
27 parts HCN or one gram KHS0 4 will liberate "1985 grams 
HCN from a solution containing an excess of cyanide of 

Supposing, however, that the cyanide of potassium solution 
contains a free alkali, the sulphuric acid is neutralized, and less 
hydrocyanic acid is liberated, in direct proportion to such alkali 
present. If there be no protective alkali, I know that in presence 
of an excess of KCN one gram of KHS0 4 liberates '1985 
grams HCN. 

If less HCN is liberated it practically amounts to its being 
retained by the alkali, so that if I take the equation — 

2HCN + CaO = Ca(CN 2 ) + H 2 0. 

54 parts of HCN=56 parts Calcium Oxide. 

The proposition is now as follows : — 

Take 500 c.c. of a cyanide solution, add one gram of KHS0 4 , 
boil for 45 minutes, and collect the HCN in a flask containing 
caustic potash, passing the vapour through a Liebig's Condenser. 

I estimate the HCN in the ordinary way in the potash solution 
with nitrate of silver. 

Let A = amount of HCN liberated by one gram KHS0 4 in 
presence of excess of cyanide without any alkali. 

Let B = amount found in experiment with alkali. 

A-B = C C x |£ or 1*037 « amount of calcium oxide 


I made a series of experiments with various proportions of care- 
fully prepared lime water and cyanide solutions, and obtained 
very accurate results — as for instance, the following. I took 300 
c.c. distilled water containing two grams of pure cyanide of potas- 
sium, and added 200 c.c. of saturated lime water (at 18° Cent.) 
which contains 0*260 grams calcium oxide. On treating the solu- 
tion as described I obtained the following results : — 

I had taken in this case 2 grams KHS0 4 . 

HCN liberated by 2 grams KHS0 4 = 0*3970 grams. 
HCN liberated in experiment = 0*1509 „ 

•2461 grams. 

C = 02461 Cx 1-037 = 0-2552 grams. 
Calcium oxide found 0-255 „ 
„ „ taken 0-260 „ 

or in percentages 0*051 and 0-052 respectively or only 10 1 00 of 
a per cent, from the correct amount, or within 2 per cent, of the 
lime taken. Of course the operation requires care and practice. 




Fob the Assay op Cyanide Solutions. 

If }- pint of 


If 4-pint of 


One Ton of Solution 



Ton of Solution 

gives Fine 

will give Fine Metal. 

gives Fino 

will j 

give Fine Metal. 

Metal : 

Metal : 


ozs. dwts. grs. 



dwts. grs. 





5 20 





8 18 





11 16 





14 14 


1 35 



17 12 


1 9 





1 145 



3 8 


1 20 



6 6 


2 1*5 



9 4 


2 7 



18 8 


4 14 



7 12 


6 21 



16 16 


9 4 



5 20 


11 11 





13 18 



4 4 


16 1 



13 8 


18 8 



2 12 


1 15 



11 16 


1 2 22 



3 8 


Fob the Assay op Cyanide Solutions. 

































































































If i lb. of Solution 

One ton will 

If |lb. of Solution 

One ton will 

gives of fine 

give of fine metal 

gives of fine 

give of fine metal 




ozs. dwts. grs. 


ozs. dwts. grs. 




14 22 




16 15 


13 25 


18 8 




1 16 16 




2 15 


1 2*5 


3 13 8 


1 7*25 


4 11 16 


1 12 


5 10 


1 16 


6 8 8 


1 20 


7 6 16 


3 16 


8 10 


5 12 


9 3 8 


7 11 


18 6 16 


9 4 


27 10 




36 13 8 


12 23 


45 16 16 

Some Useful Constants. 

The cubic content of a circular vat in feet = Dia 2 in 
x *7854 x depth in feet. 

1 cubic foot of water = 62 J lbs. nearly. 

1 ton of water contains about 36 cubic feet. 

1 gallon equals 10 lbs. 

1 lb. avd. equals 7000 grains. 
To converts lbs. troy into lbs. avoirdupois x -82286. 
To converts lbs. avoirdupois into lbs. troy x 1-2153. 





The appliances in use are of much the same nature at all cyanide 
plants, but their size, shape, and arrangement are subject to 
endless variations, being chiefly affected by local conditions, the 
character of the material to be treated, and the individual taste 
or fancy of the metallurgist. In all cases the designer should 
utilize the natural advantages at his disposal ; and, where possible, 
the solution vat, leaching vats, vacuum-cylinder, storage tanks, 
zinc extractors, and sumps should be placed on three separate 
tiers or platforms, so as to permit of the circulation of the solu- 
tions by gravitation, as shown in the following diagram : — 



A — Dissolving Tank. B — Solution Tank. 
C — Leaching Vat. D — Extraotob. E — Sump. 

When a vacuum-cylinder is used, a storage tank must be placed 
below it to receive the solution when the cylinder becomes full. 
The bottom of the storage tank must be above the extractor. 

Whatever arrangement is adopted, the solution has to be 
pumped from the sumps, either to the solution tank B, or directly 



to the leaching vat. In the latter case the cyanide solution is 
made up to the working strength in the sump, the dissolving 
tank A being placed on the floor level ; or the strength of the 
sump solution having been ascertained, the required volume is 
pumped up to the leaching tank and brought up to working 
strength by dissolving the requisite quantity of cyanide, placed, in 
a perforated box or tray, at the discharge end of the solution 

Where the height necessary for gravitation cannot be obtained, 
the need of constructing high platforms, with a correspondingly 
high building, is obviated by making up the working solution in 
one of the sumps, with each of which the dissolving tank and 
solution sump is connected. To do this without inconvenience 
an extra sump should be provided. 

This arrangement is shown in the following diagram : — 



A— Dissolving Tank. B— Sump, also used as Solution Tank. 
C — Leaching Vat. D— Extractor. 

The necessary appliances for a successful and well-equipped 
cyanide plant, where the ore is to be treated by percolation, are 
as follows : — 

1. A dissolving tank. 

2. Solution vats. 

3. Leaching or percolating vats. 

4. Vacuum-cylinder and air-pumps if percolation slows down. 

5. Storage vats. 

6. Zinc precipitation boxes. 

7. Sumps. 

8. Solution pump. 

9. Fresh-water pipes connected with every tank and vat. 

10. Line of pipes connecting vacuum-pump and cylinder with 

leaching vats. 

11. Line of pipes connecting solution pump with each sump, 

leaching vat, and solution tank. 



12. Lines of weak and strong solution pipes connecting leaching 

vats with extractors. 

13. Filter vat for washing and drying gold precipitates. 

1 4. Melting and roasting furnace. 

15. Assay office and laboratory. 

In the case of double treatment, or intermediate filling, ad- 
ditional vats are required. They are generally placed over, or at 
a higher elevation than the leaching vats. 

Dissolving Tank. — This is constructed of wood, iron, or 
steel. When made of wood, the staves are of 2 in. or 2| in. pine ; 
and when of iron or steel, the plates are from J in. to £ in. thick, 
according to the size of the vat, and stiffened with angle-iron and 
hoops. The size varies from 3 ft. to 6 ft. in diameter, and 
from 2| ft. to 4 ft. in depth. 

In large cyanide works a perforated tray, to hold the solid 
cyanide salt, is suspended over the tank by a chain or steel-wire 
rope, running over a pulley fixed to a beam overhead. The end 
of the rope or chain passes over a second pulley on the same 
beam, fixed at a sufficient distance to permit a balance-weight on 
the end of the rope to clear the side of the tank. 

In practice the solid salt is taken out of the original packing 
case and cleaned, by removing all adhering particles of sawdust 
or other packing material with a husk-broom. It is then broken 
into small pieces and placed in the perforated tray, which is 
allowed to subside into the solution. The rapid dissolution of the 
cyanide salt can be effected by imparting motion to the tray by 
pulling at the weighted end of the rope. 

The discharge-hole from the dissolving tank should always be 
placed three or four inches above the bottom, so as to allow a 
settling space for the impurities contained in the cyanide. 

The impurities contained in commercial cyanide consist princi- 
pally of black carbide of iron, and other insoluble matters, which 
would, if permitted, obstruct the solution pipes, choke up the 
filter webs, and thus cause vexatious delays. Besides this, carbide 
of iron decomposes the potassium cyanide solution of gold ; hence 
its presence in the solutions would tend to cause a loss by precipi- 
tating a portion of the gold. 

Solution Vats. — These are used for making up the cyanide 
solutions to the working strength. They are open circular tanks 
from 14 ft. to 20 ft. in diameter, and from 4 ft. to 14 ft. in 
depth. They are generally constructed of well-seasoned pine. 
The sides are made of 5 in. or 6 in. planks, 2J in. or 3 in. thick. 

The bottom is constructed of planks 12 in. by 3 in., bolted and 
dowelled together independently of the sides. The bolts are 
placed about 3 ft. apart, and are made of £ in. or 1 in. round 


iron. The washer and nut are well countersunk into the plank, 
and after being tightened, the whole is plugged up with a block 
of wood to prevent leakage around the bolt. The dowels consist of 
| in. round iron, and are 6 in. or 8 in. long. They are placed 
about 12 in. apart where the planks come near the circumference. 

The bottom is rebated into the side planks, which are kept tight 
by round iron hoops, f in. to 1^ in. in diameter, having three or 
more cast-iron turn-buckles on each hoop. One, and sometimes 
two, rings of J in. round india-rubber are placed round the 
bottom before the staves are put on. The rubber rings are 
placed in grooves which are made with a steel tool. 

The side planks, or staves, are kept in their places by the 
pressure of the hoops alone, which are generally placed from 1 5 
in. to 18 in. apart, with an extra hoop at the bottom. The hoops 
are placed, on very large deep tanks, only six or seven inches 
apart at the bottom, where the pressure is greatest, the distance 
apart gradually increasing to 18 in. or 20 in. at the top. The 
extra hoop is placed as close to the bottom hoop as the turn- 
buckles will permit. 

One or more solution vats may be required according to the 
size of the plant. 

Leaching or Percolating Vats. — These are made of many 
different shapes, sizes, and kinds of material. At first, small 
square tanks of wood were used, but the difficulty of keeping 
these tight led to the adoption of circular vats, which are also 

In Australia, the favourite material is wood, but in South 
Africa and New Zealand, steel vats are at present preferred to 
wooden ones, in all the recently erected cyanide plants. In 
America, both steel and iron vats are preferred to wooden ones. 

In Victoria and New South Wales, leaching vats made of 
ordinary corrugated iron with wooden bottoms have been used 
with very satisfactory results. Mr. W. Eddowes, who used them 
in Victoria, informed the author that they were light, strong, 
cheap and durable. For vats of large capacity, corrugated iron 
of No. 16 gauge is employed. The cost is said to be half that of 
steel vats of the same capacity. "When the value of corrugated 
iron becomes better known, it will doubtless be more largely 
employed, more especially in outlying mining camps where trans- 
port costs are heavy. 

The construction of the circular wooden leaching vats is in 
every respect the same as that of the solution vats already 
described, differing only in being provided with discharge-doors. 
The first vats in use at the cyanide works of the Waihi Gold 
and Silver Mining Company were 22 J ft. in diameter, and 4 .ft. 


deep. The sides were built of 5 in. kauri planks, 3 in. thick, 
bound together with five hoops of round iron, three of which were 
f in. in diameter, and two 1 in., having three turu-buckles on 
each hoop. Five inches is taken off the depth by the false 
bottom, filter-frame, and cloth. At the Company's new mill, 
there are ten concrete tanks in two rows, each tank being 50 ft. 
by 40 ft. and 4 ft. deep. 

The foundations or supports of wooden vats, 20 ft. in diameter, 
with 7 ft. staves, to hold a charge of 70 tons of tailings, designed 
by the author for the Moanataiari Gold Mining Company, 
Thames, consisted of mudsills, 15 in. by 6 in., laid flat, on which 
rest the sole pieces, 10 in. by 8 in., supporting the props or 
uprights, 8 in. by 8 in., on which are laid the bearers of the same 
dimensions as the sole pieces. There are five rows of props — two 
rows of three, three rows of five — twenty-one in number altogether, 
spaced with 4 ft. centres both ways. The bottom of the vat rests 
on the bearers, which are cut so as to allow a space 4 in. wide 
between their ends and the side planks. This space is left to 
detect and repair any leaks that may take place round the sides 
of the vat. 

Where the solid rock is exposed to build upon, the mudsills can 
be dispensed with. (For detailed drawings, see Plate I.) 

The same support can be used for iron or steel vats, but it is 
advisable to place 9 in. by 3 in. planks across the bearers, so as to 
more evenly distribute the weight of the vat, the bottom of which 
rests on the planks. 

At the Simmer and Jack Cyanide Works, Johannesburg, the 
leaching tanks are constructed of pine. They are 42 ft. in 
diameter, and 14 ft. deep ; bound together with fifteen hoops 
of round iron. They rest on piers of solid masonry. The staves 
and bottoms are made of 9 in. by 3 in. material. 

At the " Main Reef " Works there are six leaching vats, each 
26 ft. inside diameter, with 8 ft. staves, and holding 135 tons of 

The staves are 4£ in. wide and 3 in. thick, and planed to 
the level by machines, and afterwards hand-dressed on the abut- 
ting edges. They are checked f in. to fit on the bottom, with a 
chime 6 in. below the check. 

The bottom of the vat is made of 9 in. by 3 in. deals, planed 
by machine and grooved f in. by \ in. by a saw, and is also hand- 
dressed on the edges. Clear-pine tongues, 1 in. by f in., fill the 
grooves. The joists across the tunnel, below the vat, consist of 
9 in. by 3 in. deals, bolted together in pairs, and laid 2 ft. 3 in. 
apart from centre to centre. 

These joists are first laid in position, then the bottom of the vat 

I I 



1 ». 

N N ' 


is laid down, cramped up, and the circle struck out. The bottom 
is now sawn to the circle, and when bevelled all round is ready 
for the staves, which are driven up as tightly as possible. 

Six hoops of round iron are used to keep the staves in their 
places ; the top pair 1 in. in diameter, the middle pair 1^ in., and 
the lowest pair If in. in diameter, with screwed ends. 

Each hoop is made in three sections, rolled to the required 
curve, and connected by cast-iron turn-buckles. 


Fig. 1. — Showing Construction of Turn-buckle. 

Scale : £ in. =1 ft. 

The screwed ends pass through the turn-buckles, and while 
each hoop is being drawn up it is hammered with a heavy sledge- 

Two carpenters, practised at the work, can dress the material 
for a vat, 28 ft. in diameter and 8 ft. deep, in about a week, and 
erect it in about four days. 

With large vats, constructed of brick and cement, in an exca- 
vation in the ground, there is no means of ascertaining what leak- 
age is going on; and, in a process in which gold solutions are 
being dealt with, an exceedingly small leak, in the course of the 
year, would represent a considerable loss. For this reason their 
construction cannot be recommended where material is procurable 
for the construction of wooden or steel vats. 

At the Langlaagte Estate Company's Cyanide Works, S.A., the 
tanks are round and constructed of brick, faced with hydraulic 
cement. Their size is 40 ft. in diameter and 10 ft. deep. At the 
Crown Reef Works the tanks are also of brick, 40 ft. square and 
10 ft. deep. At the Waihi-Silverton Works, N.Z., the tanks are 
constructed of jV in. steel, being 16 ft. in diameter and 4 ft. deep. 
At the Moanataiari mine, the steel vats are 20 ft. in diameter and 
7 ft. deep. At the Cripple Creek Gold Exploration Company's 
Works, in Colorado, the vats are made of iron, being 20 ft. in 

For the direct treatment of dry-crushed ore, the leaching vats 
are seldom over 4 ft. deep, on account of the difficulty of percola- 
tion with a greater depth of ore ; but, with tailings comparatively 
free from slimes the depth varies from 8 ft. to 14 ft. 


The leaching vats, on account of the enormous weight they hold, 
must be built on strong firm foundations, so as to prevent settling, 
and the leakage which would be sure to follow. In South Africa 
they are often built on piers of stone ; and in New Zealand and 
Australia, where timber is plentiful, on massive frames of wood. 
Whatever the foundations, there should always be free access to 
the bottom of the tanks, so as to be able to detect and repair 

Each tank is provided with a separate drain-pipe, 1 J in. or 2 in. 
in diameter, with two stop-cocks near each other, one over the 
strong solution launder or pipe, the other over the weak solution 
launder, leading to their respective zinc extractors. When filtra- 
tion is assisted by a vacuum, a third stop-cock is provided for the 

The sizes of the pipes for charging the vats with the solutions 
are as follows : — 

Vats, 20 — 24 ft. in diameter, . . 2£ in. 

„ 24—32 ft. „ ... 3 in. 

„ 32 — 40 ft. „ ... 4 in. 

Two or three lines of pipe with two mains, one for the weak and 
one for the strong; and with three mains, one for the weak, 
medium and strong respectively) running parallel with the line of 
leaching vats, afford the simplest, most economical, and effective 
method of collecting the solutions as they percolate from the vats. 
This system enables the solutions from each vat to be tested separ- 
ately and readily, and by this means any mishaps can at once be 

Instead of having stop-cocks on the end of the drain-pipe, from 
each vat, a short length of rubber-hose is sometimes fixed on the 
end ; and by moving the hose the solution can be drained into the 
strong or weak launder as required. 

Steel and iron vats are now coming into general use. They 
possess many advantages over wooden ones. They are generally 
coated with a composition consisting of a mixture of coal-tar, 
pitch, and kauri gum. (See Plate II.) 

The filter-frame in steel vats is supported on a ring of iron 
riveted to the side about 3 in. from the bottom. The filter 
webbing is laid on the frame and kept in its place by means of a 
ring of angle-iron, which is constructed in four, six or eight pieces, 
or lengths, so as to be easily handled. The ordinary method of 
grouting the cloth between the ring of iron and the side of the 
vat with a small rope is the best. 

In large steel vats, the ring for the reception of the filter-cloth 
is omitted, and in this case the filter-frame is constructed so as to 

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leave an annular space an inch wide all round the vat so as to 
permit the filter-cloth to be grouted between the frame and the 

Filter-Frames. — The old filter-beds of gravel and sand have 
been entirely superseded by light wooden frames, over which are 
placed filter-cloths or webs, consisting of either extra strong 
Hessian, loose canvas, cocoa-matting, or burlap. For the filtra- 
tion of slimes, or dry-crushed ores, which always contain a large 
percentage of very fine sands, a webbing of strong Hessian or 
canvas is used; and for tailings or concentrates a webbing of 
cocoa-matting or burlap. 

In some plants a duck-cloth filter is laid over cocoa-matting. 
The duck-cloth for a 100 ton vat costs about £2, 10s. ready made. 

At the Waihi Cyanide Works the filter-frames, designed by the 
general manager, Mr. H. P. Barry, consist of narrow laths placed 
parallel, and about an inch apart. On these laths are nailed, 
transversely, narrow moulding-like laths, also about an inch apart. 
An open frame- work or grating is thus obtained, having openings 
an inch square. 

At the Main Reef Cyanide Works, Johannesburg, the filter 
frame is made of 3 in. by 1 in. slats, placed on edge, 6 in. apart, 
their ends being kept 1 in. from the sides of the vat. Strips of 
wood, 1 in. square, are nailed on the top of the slats 1 in. apart 
to form a support for the cocoa-matting. 

The filter-frames for large vats are constructed in sections. 
The sections, when fitted together, form a circular frame about an 
inch less in diameter than that of the vat. This leaves an 
annular space between the frame and the vat, which permits the 
filter-cloth to be firmly grouted in its place by means of a small 
rope passing round the circumference of the vat. 

The author has used such filter-frames for over five years, and 
finds that they possess many advantages over the old gravel- 

At plants where intermediate filling is adopted, the intermediate 
vats are provided with automatic distributors. 

Vacuum-Cylinder and Air-Pump. — Filtration is generally 
assisted by creating an artificial vacuum below the filter-bed. 
The means mostly adopted in New Zealand and Australia to pro- 
duce a vacuum is an air-tight boiler, or cylinder, connected with 
an air-pump. In large cyanide plants two air-pumps and 
cylinders and storage tanks should be provided. 

The cylinders are generally constructed of £ in. boiler-plate, 
with f in. ends. They are made of different sizes, according to 
requirements, from 6 ft. to 13 ft. in length, and from 3 ft. 6 in. 
to 6 ft. in diameter. They are provided with a solution-gauge, 


vacuum-gauge, air-cock, and man-hole, as well as the necessary 
pipe connections. 

The air-pump is single or double-acting, 7 in. in diameter, with 
an 8 in. stroke, making from 80 to 120 strokes per minute, and 
capable of producing a vacuum of 26 in. of mercury in the 
cylinder. To prevent heating of the valves it should be sur- 
rounded with a water-jacket, through which a current of cold 
water can continually circulate when the pump is running. The 
air-pump at the Waihi cyanide plant is 14 in. in diameter, and 
has a 22 in. stroke. 

All the stop-cocks, valves, pipes and connections about the 
cylinder, air, and solution pumps, and tanks, which are subject to 
cyanide solutions, should be of black iron. 

When the vacuum-cylinder becomes full, the solution is dis- 
charged into a storage vat, from which it slowly drains through 
the zinc extractor. In order to give timely warning when the 
cylinder became full, the following simple electrical contrivance 
was used by Mr. Arthur Wilson, the manager, at the Cassel Gold 
Extracting Company's tailings plant at Waihi. 

A small Erdmann float, with a platinum wire fused into the top 
and coiled into a flat helix, was placed in the solution gauge-tube. 
Two platinum wires were also fused into the upper end of the 
gauge- tube, projecting into the tube, opposite to each other, but 
not in contact. The platinum wires were connected with a small 
Leclanche battery, and when the float rose in the solution-gauge 
to the platinum wires, metallic contact was established, and an 
electric bell in the circuit sounded an alarm. 

Discharge of Leached Residues.— Where there is a 

plentiful supply of water, with a good head, the easiest and 
cheapest method of discharging the residues from the leaching 
tanks is to sluice them out by a side-door. At the cyanide plant 
of the Waihi Gold Mining Company the residues are sluiced out 
by two 2 in. hose-pipes under a head of 150 ft., giving a pressure 
of 65 lbs. to the square inch. 

At the Witwatersrand Goldfields, where there is a scarcity of 
water, and often a want of fall for the sludge, the " bottom dis- 
charge " is largely practised, the residues being shovelled through 
a hole in the bottom of the vat into a truck immediately below. 
At the Barret Company's Works the tailings are shovelled into a 
launder below the vat, and a stream of water carries them away. 

At the Crown, Woodstock and Talisman Cyanide Works in 
New Zealand, where the ore contains a percentage of coarse gold, 
the residues are slowly sluiced over extensive amalgamated copper 
plates placed immediately below the discharge holes. 

At the Langlaagte Company's Cyanide Works, near Johannes- 


burg, the residues are discharged from the large brick leaching 
tanks by means of steam travelling cranes, which lower the bodies 
of the empty trucks into the tanks, where they are filled by Kaffir 
labour. When filled the trucks are raised and placed on their 
carriages, to be wheeled away to the dump. 

Discharge Doors. — When side-discharge by sluicing is used, 
one or two outlets are generally provided for each vat ; but in the 
case of bottom-discharge there are two, four, six, or eight dis- 
charge openings to each vat, according to its size. 

Fig. 2. — Butters' Bottom-Discharge Door Scale : J in. =1 ft. 


At the Witwatersrand Goldfields the bottom-discharge 
employed for discharging the round wooden leaching vats. When 
filling a deep tank with tailings, a length of wrought-iron pipe or 
cylinder, three or four feet long, is placed over each discharge- 
hole, and then the tailings are dumped in. The pipe raises the 
outlet within a few feet of the surface, and thus facilitates the 

On these fields, Butters' bottom-discharge doors are largely 
used. Fig. 2 shows their construction. 

The arrangement is very simple and effective. On the bottom 
side of the tank, a cast-iron ring, A, is bolted to the cast-iron 


cylinder, B, inside the tank. Inside the cylinder is a projecting 
lug, C, upon which rests the hanger, D, which forms part of the 
screw, E. The cast-iron cover, F, when placed in position, is 
fastened by the butterfly-nut, G, and by screwing this firmly the 
whole arrangement becomes watertight. The faces of the ring 
and cover should be planed perfectly even, so as to make a good 
joint. The joint is also made tight by a luting of clay. 

Fig. 3 shows another bottom -discharge door, designed by Mr. 
W. F. Irvine. It is simple in construction and not likely to get 
out of order, but would be more convenient in shallow than deep 

Plate II*. represents the working drawing of an improved 
bottom-discharge door, simple in construction, very efficient, and 
easy to handle. It was originally designed by Mr. G. R. Walker, 

Fig. 3. — Irvine's Bottom-Discharge Door. 
A — Recesa for Packing. S<ale : J in. = 4 in. 

' improved by Mr. A. Price, and perfected and used in its present 
form by the author in the Moanataiari and other cyanide works. 
The Koppel patent tank door recently placed in the market differs 
from the above only in that the pressure to close the door is 
applied by a screw at the side instead of at the centre. 

The door is swung on a loose hinge, and faced with a wide ring 
of rubber-insertion. It is closed by lifting, or swinging the door 
with the left hand over the discharge hole, and then turning 
an iron button with the right hand so as to carry the weight of 
the door, which is now approximately in position. The loose dog 
is then inserted in the lugs and the door screwed up tightly with 
both hands. The iron-button turns easily on a small stud-bolt, 
and serves to keep the door in position preparatory to the final 



To open the door the dog is unscrewed, and the button turned 
to one side, when the door falls open by its own weight, swinging 
away from the workman. The operation of opening or closing 
occupies two or three minutes. 

The drawing shows a door designed for a steel vat, but it could 
easily be adapted for a wooden vat by lengthening the cylindrical 
part, which projects into the vat the necessary length to make it 
level with the filter-frame, less, of course, the thickness of the 
coir matting or filter-cloth. 

Plate X. in Chapter XI. shows the bottom discharge-door 
designed by Mr. Roche for the Waihi Company's new cyanide 
plant at Waikino. 

Fig. 4. — Side-Discharge Door. Scale : f in. = 2 ft. 

Fig. 4 shows a modification of a side-discharge door for wooden 
vats, designed by Mr. W. R. Feldtmann, of Johannesburg. 

Another method of side-discharge, designed by Mr. Irvine, is in 
use at the Crown Reef Company's fine cyanide works, where the 
large square brick and cement vats are provided with doors which 
permit of the ingress of the discharging trucks.* The door frames 
are bolted to the cement walls, and the plate-iron doors are drawn 
tight against these by means of an ingenious arrangement of 
sliding lugs, bolts, and nuts. 

The side-discharge sludge doors used in New Zealand and 
Australia are of the simplest construction and yet perfectly effi- 
cient. They consist of a cast-iron frame with two projecting lugs, 
one on each side, and a projecting bar running along the top side. 

* Feldtmann, Notes on Gold Extraction, 1894, p. 3. 


The lug on the right side is placed with the notch upwards, that 
on the left with the notch in the reverse position. 

The opening is closed by a cast-iron door, which is kept in 
position by the pressure of a screw acting through a loose iron 
dog, the ends of which fit into the lugs so as to obtain the 
necessary leverage. 

The door is suspended in front of the opening, preparatory to 
fixing up, by a hook of bent round iron, which is supported on 
the projecting bar on the frame. It is rendered watertight by a 
facing of rubber insertion, fixed on with tar, or by a luting of clay. 

These doors seldom give any trouble. They are easily opened 
or closed by a few turns of the screw. The different parts are 
shown to scale on Plate III. 

Steel leaching tanks with bottom-discharge, designed by the 
author for the Moanataiari Cyanide Plant, are shown on Plate II. 

Sumps. — There are at least two of these in every cyanide 
plant, to receive the cyanide solutions after passing through the 
zinc-extractor, one for the strong solution and one for the weak. 
In plants dealing with acid ores or tailings there is often an 
additional tank besides these for storing the alkaline wash- 

The size of the storage sumps depends on the size of the plant. 
In most plants they are the same size as the leaching vats. They 
are constructed either of steel, concrete, bricks faced with cement, 
or wood. The latter is the favourite material. The construction 
of wooden sumps is the same as that of the storage tanks or vats. 

In many plants the sumps are placed below the level of the 
lower or extractor floor of the cyanide building, and in such cases 
they are decked over with planks, having a man-hole for repairs 
or cleaning. The depth of the solution is indicated by the 

In some tailings plants the working cyanide solution is made 
up in strong solution sumps, thus saving the construction of a 
storage vat, but requiring an extra sump. This method could not 
be used with advantage for the treatment of dry-crushed ores in 
which it is necessary to apply the strong solution slowly from 
below, so as to prevent the formation of lumps and channels in 
the pulp such as would occur were the solution turned on to the 
dry ore from above. 

Zinc Extractors. — There are at least two of these in every 
plant, one for the strong solutions and one for the weak. In 
works treating sands and slimes it is usual to provide a set of 
three extractors, namely for the strong, medium and weak 
solutions respectively. They are constructed of wood, and 
consist of oblong boxes, each divided into a number of compart- 


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ments, generally eight, ten, or twelve, by means of partitions and 
baffle-boards fixed in such a way that the solution finds its way 
upwards through the zinc-turnings in each compartment. 

The first and last compartments are filled with sand or tow to 
act as filters, the upper one to remove any clayey matter in the 
solutions, the lower one to prevent the escape of slimes. 

The remaining divisions are sometimes provided with removable 
shallow trays of wood, with convenient handles, often of bent 
round iron, to enable them to be easily lifted up when necessary. 
The trays support, at the bottom, a wire screen of £ in. mesh, 
which permits the free circulation of the solutions, and enables the 
gold and zinc slimes, as they form, to fall through into the 
bottom of the compartment. 

Instead of projecting handles, the trays are often provided 
with an iron plate on two sides with two holes in it to take the 
end 8 of hooks for lifting the trays and their contained zinc. 

The extractors are cleaned-up from plug-holes on the side, one 
to each compartment. To facilitate the " clean-up " the bottoms 
of the compartments slope to the side, and also to the lower side- 
corner, where the discharge-hole is situated. 

A launder of wood or iron is fixed on the discharge-side, 
immediately under the plug-holes, so as to receive the slimes when 
they are washed out of the boxes. The top of the extractor, as 
well as the side launder, is protected with a lid or close-fitting 
grating of wood or iron, provided with locks, to prevent the con- 
tents of the extractor being tampered with. 

To facilitate the withdrawal of the wooden plugs from the 
discharge-holes, short lengths of rubber-hose are fixed in the holes. 
The rubber is yielding, and, while rendering the holes water-tight, 
enables the plugs to be withdrawn without the force which is 
generally necessary when they are driven into the bar-holes. 

The extractor should be constructed of well-seasoned pine, with 
sides of 1 J in. or 2 in. boards, and the divisions of 1 in. boards, 
well dressed. The size will depend upon the capacity of the 
plant, and will vary from 12 ft. to 20 ft. in length, from 2 ft. to 
3 ft. in depth, and from 15 in. to 45 in. in width. The extractors 
mostly used are shown on Plates IV. and IYa. 

It is now well known that the bulk of the precipitation takes 
place in the first two or three boxes, and in cyanide plants hand- 
ling large volumes of cyanide solution daily, the tendency has 
been of late to reduce the number of divisions in the extractor 
to six or eight, and to increase their dimensions. In this case the 
extractor is simply a long box or trough divided by baffle-boards 
into compartments, without side plugs or side launder. At the 
periodical " clean-up " each compartment is cleaned separately ; or 


sometimes two or more, often all the compartments, are made to 
communicate with each other by plugs at the bottom of baffle- 
boards so that the slimes can be washed into the lower one of 
the set thus communicating, and thence be baled out into a tub 
or run into a bucket. (See Plate IVa.) 

The construction of the extractor permits an endless variety of 
design according to the individual fancy of the engineer, always 
bearing in mind that the object to be attained is to compel the 
solutions to pass upward through the zinc at a uniform and slow 

At the Waihi Company's large mill at Waikino, the solutions 
are collected in a rectangular tank divided into three compartments 
each forming a coir-filter or clarifying box, for the strong, 
medium and weak solutions respectively. 

Each compartment is provided with a false-bottom, 9 ins. 
deep, covered with cocoanut matting which is kept in place by a 
light wooden frame. 

From the clarifying boxes or compartments, the solutions run 
into the different extractor boxes, which are made 23 ft. long by 
3 ft. wide by 2 ft. 6 ins. deep. The extractors have flat bottoms 
and are divided into nine compartments, namely seven for precipi- 
tation and two for settling, one at each end, which is filled with 

The trays are 6 in. deep and covered with £ in. mesh screens. 
The slimes are washed into the end compartment through plugged 
holes in the baffle boards. 

The solutions leave each box by a 3 in. pipe connecting with 
the main leading to the sumps. The cement floor is constructed 
so as to convey all drainage from the extractors to a small well. 

The Roasting Furnace. — In Australia and New Zealand 
in small cyanide plants this consists of a shallow cast-iron pan or 
plate, built over a small furnace. A funnel-shaped hood of light 
wrought-iron is placed over the furnace to carry off* the zinc oxide 
and other fumes. 

The hood is suspended by a chain or steel-wire rope passing over 
a pulley overhead, and is balanced by a weight on the end of the 
chain. The upper or narrow end of the hood telescopes for a few 
feet into the pipe leading to the condensing flue, and by means of 
the balance weight the hood can be lowered or raised over the 
roasting pan or plate as required. 

The first part of the condensing flue is a nearly horizontal 
length of iron-pipe, on which a stream of cold water is sometimes 
allowed to play, so as to condense the mercury vapours and zinc 

In large cyanide works, a built-in muffle-like roasting furnace 

Plate IV 


constructed of fire-brick is commonly used. It possesses the 
advantages that a better regulation of the heat can be effected, 
and that with care the loss through dusting is les3 than with 
open pans. 

A roasting furnace capable of taking 50 lbs. of precipitate at a 
time is shown in Plate V., figs. 1 and 2. 

Fig. 3, Plate V., shows a special furnace for roasting slimes. It 
is a com oined reverberatory and muffle furnace. There is a fire- 
box at one end, a cast-iron hearth in the middle, on which the 
material lies, and an exit-flue at the other end. The fire-bridge is 
built in two parts, with a passage between them leading to flues 
below the cast-iron plate, so that part of the heat is deflected and 
passes under the plate, while the remainder passes over it with a 
plentiful supply of air, thus effecting the oxidation of nearly all 
the zinc. 

The Filter-Press or Vat. — In large works, where hundreds 
of pounds of precipitates have to be handled in the course of a 
week, a filter-vat with a strong close canvas webbing is found very 
convenient for washing and drying. The false bottom of the vat 
is connected with the air-pump. In plants where the slimes are 
acid treated, a small filter-press is used for washing and drying. 

Tell-Tales. — These should be provided for the dissolving and 
solution tanks, and the storage sumps. They should be marked 
in feet and inches, so that the workmen can draw off a given 
depth of solution without any calculation. The cubic content of 
an inch of depth of each vat or tank in pounds should be marked 
on a board, so that when so many pounds of solution are required, 
their equivalent in inches of depth can be ascertained at a glance, 
thus saving delay and error from needless calculations. 

Solution Pump. — This is generally a small centrifugal pump, 
with a 2 in., 3 in. or 4 in. discharge-pipe. It is driven off the 
shaft which conveys power to the air-pump. 

The Air- Lift for Sands, Slimes, Solutions, etc.— On 

sloping ground the milling plant and cyanide works can generally 
be so arranged that the sands and slimes find their way to the 
leaching vats by gravitation. In cases where natural slopes are 
not available for mill sites, the battery sands have to be lifted to 
the leaching vats by artificial means. In South Africa, bucket- 
wheels; in America, bucket- wheels and sand-pumps; and in 
Western Australia, bucket-wheels sand-pumps and air-lifts, are 
commonly employed to lift the tailings. 

The air-lift was first introduced and worked on a large scale, 
some four years ago, by Mr. J. W. Archibald, manager of the 
Mount Malcolm Proprietory Gold Mines, Limited, at Murrin- 
Murrin, in Western Australia. Since then it has come into 


favour in Kalgoorlie, and seems likely in time to supersede both 
tailings wheels and sand-pumps in Western Australia. 

So far the air-lift has not shown a high efficiency for the 
power employed ; but it possesses several special advantages which 
seem to recommend it to mine owners, namely : cheapness of 
erection, continuous operation, and little wear and tear, having 
no complicated parts to get out of order. The Associated Mine 
has two in operation ; Kalgoorlie, six ; Boulder No. 1, four ; 
Westralia Mount Morgans, two ; Lake View Consols, one ; besides 
many others. 

Judging from the success which has attended the use of air- 
lifts in Western Australia, it is certain that this form of elevator 
is deserving of more general application, and for that reason 
I append the following lucid description of their construction, for 
which I am indebted to the courtesy of Mr. Max von Bernewitz 
of Kalgoorlie. 

He says : " In erecting an air-lift the first work is that of sinking 
a well-hole the same depth, or better still, 2 ft. more than the 
height the pulp has to be lifted. The size of the hole depends 
on the size and number of lifts to be erected : for example, if one 
6 in. lift is to be put up, a well-hole not less than 3 ft. in diameter 
should be sunk. The reason why a hole is sunk is that the lift has 
always in it a column of water and pulp, equal to height to be 
lifted, whereby the air meets with the necessary resistance required 
to start the column in motion. 

" Next, a large pipe, say 2 ft. 6 in. in diameter, with a tapered 
bottom and of the same depth as. the well-hole, is lowered into the 
latter. The delivery pipe or lift is then fitted together for the 
correct length, and lowered into the large pipe. Finally, the air- 
pipe is fitted together and run down at the side of the lift-pipe. 
At the bottom of the air-pipe is a bend which looks up into the 
lift about 2 ins. It is desirable to have flanged joints on all the 
pipes, in order that they may be quickly and easily disconnected 
in case of an accident. 

" At the top of the lift-pipe, a bend may be fitted on, but the 
practice now is to bolt a launder directly on to a flange at the top 
of the pipe. By this means the air, as it reaches the top, can 
escape at once." 

It has also been found that the larger the air- and lift-pipes, 
the better is efficiency attained ; also, that air-lifts should be in 
duplicate in case of mishap. 

If the air is taken from a receiver at, say, five atmospheres 
pressure, a pressure several times greater than actually required by 
a lift, and expanded before doing its work, it is found that there is 
a loss of power. But if the air is taken direct from the compressor 


-18 — —18- 

« . 





to the air-lift, a far greater efficiency is obtained, as the rising and 
falling of the level of the pulp acts as a perfect governor to the 
compressor. The practice on most of the mines is to take air 
from a receiver At, say, five atmospheres pressure, and conduct it to 
a lift through a reducing valve ; but this is unsatisfactory, and a 
separate compressor giving large quantities of air at a low pres- 
sure is the more economical method. 

Again, when a compressor supplies air for many purposes the 
working of the lift is greatly interfered with. Take, for example, 
the mill at the Kalgoorlie Mine. Here a compressor running at a 
high speed supplies air to actuate two solution pumps ; also air to 
agitate slimes ; to force the agitated slimes into filter-presses ; to 
dry the presses and work three air-lifts, of which one lifts pulp 
20 feet, one sands 30 feet, and one clear water, 15 feet high. 
All the above require air at different pressures, and interfere 
with the working of the air-lifts. It is certain that a small 
compressor for the lifts alone would be more satisfactory. 

On the W. A. goldfields, where fuel, water, and wage costs are 
high, the working of air-lifts is fairly expensive, but in countries 
where compressors can be driven by water-power with little or 
no attendance, the cost should not be high. 

Tn the Kalgoorlie Mill a little trouble was experienced with a 20 
ft. lift, which elevates red hot roasted ore previously mixed with 
dense salt water. This lift elevates 700 tons of pulp per diem 
with an air pressure of 0*7 atmosphere. The pulp when raised 
is almost at boiling point, and a hard deposit quickly sets around 
the pipes, and gradually chokes them up at the bottom. This 
is due to the deposition of lime and magnesia from the ore. 

In this mill a proportion of six parts of water to one part of ore 
was found necessary. The above lift-pipe was 8 in. in dia- 
meter, and air-pipe 1£ in. 

In the Hainault Mill, a 25 ft. lift has worked for over twelve 
months with no stoppages, whereas the sand-pump was always 
hung up for repairs. 

At the Associated and Westralia Mount Morgan's mills, lifts are 
in operation elevating thick slimes from settlers to the agitators. 
At the latter mill there is a lift which raises pulp from 30 head of 
stamps, crushing 120 tons per diem. Its dimensions are : — 

Height of lift-pipe above well, about 
Depth „ „ below top of well 
Total length of lift-pipe 
Diameter of ,, . 

„ air-pipe . 

Air-pressure .... 

20 feet. 
22 „ 
42 „ 
6 inches. 

0*8 atmosphere. 



/At Kalgoorlie, air-lifts are employed to circulate cyanide solu- 
tion from the bottom to the top of leaching vats, and to raise 
solutions from sumps. At the Boulder No. 1 Mill, two air-lifts 
are for lifting mercury in connection with the Reicken process, 
having the following dimensions : — 

Height of lift-pipe, about . 
Depth of well .... 
Diameter .of pipe forming well 
„ lift-pipe . 

12 feet. 
4 „ 
3 inches 
1 inch. 

„ air-pipe . 

2 » 

Air-pressure, 3 to 5 atmospheres. 

The size of the plant will depend on the intended capacity per 
month and the condition of the material to be treated. A plant 
capable of treating 2000 tons of freely-percolating tailings per 
month would probably not be able to treat more than half that 
quantity of dry-crushed ore or slimy tailings. 

With dry-crushed ores the depth of the charge seldom exceeds 
4 ft., while with sharp tailings the depth sometimes reaches 10 ft. 
or 12 ft. Cyanide plants for the treatment of dry-crushed ore are 
therefore provided with a large number of shallow vats ; and those 
for tailings with a smaller number of deep vats. 

The following gives the number of leaching vats required for 
the treatment of different quantities of dry-crushed ore and 

For Dry Orb or Slimy Tailings. — Size of vat : 22J ft. diameter, 
4 ft. deep ; charge 30 tons. 

2 vats . 

350 tons per month. 

4 „ ... 

. 700 „ 

6 , 

• 1050 „ 

8 „ . . . . 

. 1400 „ 

10 „ ... 

. 1750 „ 

12 „ ... 

• 2100 „ 

For Freely-Percolating Tailings. — Depth of vat, 10 ft. ; 
charge 100 tons. 

2 vats . . . . . 1200 tons per month. 

4 2400 

. 3600 

. 4800 






In small plants it is advisable to have a spare leaching vat in 
case of a temporary breakdown in one of the vats in use. 



The cost of the cyanide plant depends largely on the locality, 
the material used in the construction, and, to a certain extent, on 
the condition of the material to be treated. For example, a 
plant to treat 2000 tons of dry-crushed ore per month will cost 
more than a plant to treat 2000 tons of tailings. 

The approximate cost of plants of different capacities in Australia 
and New Zealand, when wood is the material used in the con- 
struction of the vats, foundations, and buildings, is given below : — 

For Dry-crushed Ore — 

Monthly Capacity. 


350 tons . 


700 „ . . . . 


1050 „ . 


1400 „ . 


1700 „ 


2000 .„ ... 


For Tailings — 

Monthly Capacity. 


1200 tons 


2400 „ . 


3600 „ . 


4800 , 


The above estimates include the cost of the laboratory, assay, 
and melting furnaces, and all the appliances for a successful 
cyanide plant, but does not include the cost of a bucket tailings 
wheel or sand-pump for elevating sands, etc. 

The cost of steel or iron vats is about the same as that of wood. 
The steel leaching vats at the Waihi-Silverton, 16 ft. diameter 
and 4 ft. deep, with central discharge, cost £56 each, and the 
wood foundations about £10 each. 

At Johannesburg the cost of a cyanide plant is about 25s. per 
ton of tailings to be treated per month. Thus a plant to treat 
3000 tons of tailings per month would cost about £4000, not in- 
cluding the cost of tailings wheel or assay and smelting offices. 
For larger plants the cost is smaller in proportion. The Hand 
cyanide plants, it should be noted, are not roofed over as they 
are in many other countries. 

In America steel leaching vats cost about 3£ cents per lb. free 


on board oars. Thus a 100 ton vat weighing 9400 lbs. would 
cost about £67. For red wood vats the cost is as follows : — 

100 tons £42 F.O.B. 

200 „ . . . . . 80 „ 
500 „ 120 „ 


The American red wood vats are stated to give every satisfac- 
tion. They are generally coated with P. and B. Paraffin paint, of 
which 40 gals., costing £8, is sufficient to coat a 500 ton vat twice 
(that is the inside only), two coats on the sides and three on the 


Synopsis Of the Process. — No hard and fast rules can be 
followed in the working of the cyanide process, as modifications 
have to be introduced to meet the special requirements of the 
different classes of ore ; but the essential features of the process, 
whether for the treatment of ores or tailings, are practically the 
same in all cyanide plants. 

The first operation is the filling of the vats. In the direct 
treatment of dry-crushed ore, the pulverized material is filled in 
to a depth of three or more feet, according to its fineness. With 
a uniform granular product, five or even six feet of ore may be 
filled into the vat, and, in the case of freely-percolating tailings, 
the vats may be safely charged to a depth of from six to ten feet. 

With tailings it is the practice to fill the vat within 20 inches 
of the top, because the strong solution or preliminary wash-water 
when applied causes the charge to shrink several inches. 

In the case of acid ores or tailings, a preliminary treatment 
with a caustic alkali, or at least a water-wash, is necessary in 
order to remove, or neutralize, the mineral salts and acids which 
decompose the cyanide solutions. 

Oxidizing agents such as barium peroxide, potassium ferri- 
cyanide, sodium peroxide, and cyanogen bromide, have been tried 
in America, South Africa, and New Zealand. The results obtained 
on a working scale with the first and last of these are very satis- 

The strong solution, containing generally from 0'2 per cent, to 
0*5 per cent, of potassium cyanide, is then applied. The proper 
strength of solution must be determined experimentally, and with 
most ores it will be found advisable to use a larger bulk of solu- 
tion rather than to unduly increase the strength. Silver ores re- 
quire a larger bulk of solution than gold ores in order to obtain a 
fair extraction. 



In the case of a dry-crushed ore the solution is allowed to per- 
meate the mass from below, but iu the case of tailings it is allowed 
to flow in over the surface, and soak downwards. The filling of 
the strong solution generally takes from two to four hours. With 
tailings it is allowed to stand iu contact for twelve hours, after 
which percolation is begun. With dry-crushed ores slow artificial 
percolation is commenced at once. 

The leaching with the strong solution takes from twenty-four 
to forty-eight or more hours, according to the character of the 
bullion ; but in the case of ores containing a proportion of fairly 
coarse gold, it is customary to make the strong solution up to the 
working strength and pass it through the charge till an adequate 
extraction is obtained. 

The weak cyanide solution, often called the first wash, is then 
applied. It is pumped from the strong sump, and generally 
varies in strength from 1 per cent, to 0*25 per cent, of potassium 
cyanide. It is allowed to percolate as rapidly as possible, the fil- 
tration being facilitated, when necessary, by means of an artificial 
vacuum, which may be created by a steam exhaust, or an air- 
pump connected with a vacuum-cylinder. 

The first weak cyanide wash is succeeded by two or three 
washes of solution from the weak sump, containing from 0*02 per 
cent, to 0*1 per cent, of KCy. 

A final washing of clean water is then applied, which serves 
to displace the preceding weak cyanide wash. By this means the 
quantity of cyanide solution in circulation remains much about 
the same. 

In the treatment of pyritic ores it is sometimes advantageous 
to apply a weak cyanide solution before the strong solution. By 
this means a larger extraction and a saving of cyanide is effected. 

The quantity of the strong solution used is about one-third of 
the weight of the ore. The cyauide and water-washes are each 
about one-sixth of the weight of the ore. 

During the treatment of slimy tailings it is found advantageous 
to sometimes turn the material over by hand-labour, and thereby 
effect a more complete washing out of the cyanide solutions from 
the slimes, which always have a tendency to entangle and retain 
them or even to transfer it from one vat to another and subject to 
a second leaching. 

Discrepancies between the theoretical and actual extractions 
are often a source of much annoyance and perplexity to the metal- 
lurgist, but in most cases they will be found to be due either to 
imperfect sampling or incorrect tonnage. At most mines a ton 
weighs considerably over 2240 lbs. 

At the Witwatersrand Goldfields, in South Africa, the tailings 


are sometimes subjected to a double treatment — i.e., after the first 
cyanide treatment the residues are charged into other leaching 
vats, and again treated with cyanide. A higher extraction will 
result, but it is doubtful if the difference would cover the cost of 
the extra handling and cyanide. 

The average extraction on the Rand where only high-priced 
white labour was available, amounts to about 72 per cent., and 
with double treatment this is raised to 85 per cent. With tailings 
of fair value the extra recovery is, however, said to leave a margin 
of profit. 

The sequence in which the different operations are applied may 
be tabulated as follows : — 

1. Filling the leaching vats. 

2. The preliminary treatment with water or alkaline washes if 


3. Leaching with the strong solution containing from 0*3 per 

cent, to 0*6 per cent, of KCy. 

4. First wash, with cyanide from strong sump, containing from 

0*1 per cent, to 0'26 per cent. KCy. 

5. Second wash, with cyanide from weak sump, containing from 

0*02 per cent, to 0'1 per cent. KCy. 

6. Third wash, same as second. 

7. Fourth wash, same as last. 

8. Fifth wash, with clean water. 

The gold and silver in the ore are dissolved by the strong solu- 
tion, and removed or carried out by the first and second washes. 
The cyanide solutions are allowed to flow through the zinc extract- 
ors, the strong solutions through the strong box, and the weak 
solutions through the weak box. The first two washes, which 
generally carry most of the dissolved bullion, are conducted 
through the strong extractor. 


Filling the Vats with Dry-Crushed Ore. — In the case 

of dry-crushed ores, the charging of the vats is a simple operation, 
the only disadvantage being the clouds of dust, which seem to be 
inseparable from the handling of dry-pulverized ores. At Waihi 
90-8tamp mill the vats were generally charged by trucks running 
on to a traveller, provided with hand-traversing gear so as to enable 
the sand to be tipped in different parts of the vat. In order to 
prevent packing, the sand was discharged from the trucks on to 
the platform of a small traveller fixed below the main traveller. 
By this means the pulverized material was dispersed in a gentle 
shower over the whole vat. 


At some plants the dry pulp is conveyed from the mill to the 
vats by screw-conveyors or by belt-conveyors, both very efficient 
methods of transport. 

At the Kapai- Vermont Cyanide Works, Kuaotunu, N.Z., the 
vats were filled directly from the dust-bin, which was situated 
overhead in an elevated position. By means of a chute, provided 
with a universal canvas joint, the material was evenly spread over 
the vats. 

When the vat was charged the surface of the ore was made 
smooth by passing a wide, shallow wooden hoe or rake over it. 

This method of filling possesses many advantages over hand- 
filling by trucks. It is cheaper, more expeditious, and healthier 
for the workmen, as it raises less dust. 

Filling the Vats with Tailings. — One of the most serious 
disadvantages of wet-crushing with the stamp-mill is the produc- 
tion of fine slimes. All ores, even the most silicious, form a pro- 
portion of slimes when wet-stamped ; and when clayey or earthy 
matter, country-rock, iron or manganese oxides are associated with 
the ore, the proportion of slimes is often excessive. 

In many places the slimes are very valuable, sometimes even 
more so than the sands. In leaching processes they prove very 
refractory, as they seriously interfere with the percolation and 
washing, thereby causing the extraction to be both costly and 
imperfect. When the slimes are irregularly distributed through 
the sands the cyanide solutions form channels through them, and 
imperfect leaching is the natural result. 

At the Witwatersrand Goldfields, where wet-crushing, followed 
by copper-plate amalgamation, is at present universal, there are 
two methods in use for dealing with the tailings before treatment 
with cyanide. The pulp from the copper-plates is lifted by a 
bucket wheel, run through a launder, and then classified by vpitz- 
kasten or spitzluten into two products, namely: — 
(a.) Sands and slimes, 80 per cent. 
(b.) Concentrates containing some sands, 20 per cent. 

The concentrates are collected in a storage tank, and kept under 
water to prevent oxidation, while the sands and slimes pass on to 
the cyanide vats for treatment, either by what is known as the 
" Direct Filling " process or by the " Intermediate Filling " process. 

At the extensive works of the Langlaagte Estate Gold Mining 
Company, the tailings, after leaving the plates, are concentrated, 
and then run into three large settling dams, each capable of hold- 
ing 7000 tons. The sands settle in the dam, while the slimes are 
carried off by the overflow, and allowed to run away. The tailings, 
now free from slimes, are hauled from the dams in trucks by 
means of two endless steel-wire ropes, and run up to an overhead 


tram-line, from which they are dumped into the leaching vats 
ready for treatment. 

Intermediate Filling 1 . — At the Robinson, Princess, and many 
other mines at Johannesburg, the tailings, after being subjected 
to concentration, are run into intermediate settling-tanks, which 
are circular in shape, and sufficiently large and numerous for the 
requirements of the mill. At the Robinson plant, treating 330 
tons per day, there are six circular wooden vats, each 24 ft. in 
diameter, and 11 ft. in depth. This gives a settling surface of 
32 square feet per stamp. 

From these intermediate settling vats the tailings are distri- 
buted to the leaching vats. When the slope of the ground 
permits it, the settling vats are placed above the leaching vats, 
the tailings being discharged from holes in the bottom ; but when 
they are below the level of the leaching vats, the tailings are 
hauled up in trucks actuated by endless wire ropes. 

To ensure a fairly even distribution of the sand and slimes in 
the leaching vats, a simple and ingenious automatic distributor, 
invented by Messrs. Butters & Mein, is now in use in all parts of 
the world. It consists of a central casting, with a vertical spindle 
revolving in a foot-step, and carrying a conical hopper at the top, 
from which radiate twelve or sixteen iron pipes with bent ends. 
These pipes vary in size from 1£ in. to 2£ in. in diameter. To 
their discharge ends are attached flattened nozzles to assist in 
spreading the tailings over a wide area. A coarse screen is placed 
over the central hopper, or bowl, so as to prevent stones or pieces 
of wood, or other debris, from choking the pipes. The distributor 
is fixed on an iron column in the centre of the vat, and the 
reaction of the pulp escaping from the bent pipes causes it to 
slowly revolve in a manner similar to that of an ordinary garden 
sprinkler. (Fig. 4, Plate V., and Plate VI.) 

To insure success, the collecting vat must be filled with clean 
water before admitting the pulp, otherwise the slimes would 
settle with the sand until the overflow of water began. While 
the machine is running, there must be a continual overflow from 
the vat to carry off the fine slimes. The discharge, or overflow, 
from the settling vat takes place at all points of the circum- 
ference, being received into an annular ring, surrounding the top 
of the vat, which conveys it to the slime-pit. The overflow of 
slimes must be continuous until the vat is full of sand, there- 
fore, when the battery stops, a regular quantity of water should 
still be supplied to the vat. 

The vats are provided with filters, and when full of tailings the 
water is allowed to drain off for fifteen or twenty-four hours, while 
six hours before discharging, holes are dug down to the discharge 


doors to let more water flow out. When drained, the tailings are 
discharged into the leaching vat below, through "bottom-dis- 
charge " doors, or into trucks, which are then run into the tank 
house, where their contents are emptied into the leaching vats. 
At the same time the required amount of lime (quicklime ground 
in a ball-mill) is added in such a way that it gets distributed 
through the whole mass. 

The advantages claimed for intermediate filling are : — 
(a.) By means of the distributor, nearly all the sands are 
collected in the intermediate tanks, the bulk of the slimes 
escaping. (The quantity of slimes depends on nature of 
ore, size of screen, and size of battery-box. ) 
(6.) When discharging through the bottom, the sands during 
the operation get thoroughly mixed, thus being in the 
best condition for treatment, 
(c.) Oxidation of pyrites is very slight, so that little cyanide 

will be consumed. 
(d.) A higher extraction can be obtained owing to the presence 
of the fine sands, from which 85 per cent, of the gold 
contents can be extracted by cyanide, 
(e.) The cost of treatment is very small. 

The principal disadvantage is the tendency of the distributor to 
form deposits of slime, through irregular distribution, at places 
on, or near, the bottom of the tank, thereby causing some trouble 
in draining off the water. 

Direct Filling. — This method is in use at the works of the 
City and Suburban, Crown Reef, and other companies at the 
Witwatersrand, and consists in conducting the pulp, as it leaves 
the copper-plates, into a classifier or spitzlute. In this the pulp is 
divided into two streams : oue, the overflow, carrying slimes and 
fine sands ; the other, carrying the coarse sands, together with 
some fine sand and slimes, which are conveyed by an india-rubber 
hose to the leaching vat, where they are distributed by moving 
the hose over the whole area of the vat. The excess of water is 
carried off by adjustable gates fitted inside the vat, taking with it 
some fine sand and slimes. 

The advantages claimed for direct filling are : — 

1. That pyritic tailings are exposed to the minimum of 


2. A second handling of the tailings is avoided. 

3. A rough preliminary classification is effected, thus separat- 

ing the fine slimes. 
The principal disadvantages of this method are : — 
1. The packing of the tailings, which prevents a complete 

draining of the contained water. 



3 ' 


2. The unequal distribution of the sands and slimes, which 

militates against a perfect extraction by favouring the 

formation of channels in the mass during the subsequent 


The method of classifying the pulp by a spitzlute, and direct 

filling, has been in operation at the Great Mercury Cyanide 

Works, Kuaotunu, N.Z., since 1892. When the leaching vat is 

charged, the water is drawn off through the filter-bed, and after 

the tailings have drained they are turned over by hand so as to 

loosen them, and at the same time thoroughly mix any slimes 

present with the sands. 


The Preliminary Water or Alkaline Wash. — This 

treatment is only necessary in the case of ores or tailings contain- 
ing the decomposition products of pyrites, or other base sulphides. 

The drying of a pyritic ore in a kiln for dry-crushing and direct 
treatment generally results in the formation of soluble sulphates, 
which are destructive to cyanide. At the works of the Woodstock 
Gold Mining Company, at Karangahake, N.Z., the consumption 
of cyanide from this cause was very heavy ; but the preliminary 
treatment with a caustic alkali, according to the later report of 
the management, has effected a great saving of cyanide, accom- 
panied by a higher extraction. 

The products of the partial oxidation of iron pyrites in tailings 
are principally free sulphuric acid, soluble sulphates, and insoluble 
basic sulphates, all of which cause a large consumption of cyanide, 
and must, therefore, be removed or neutralized before cyanide 
treatment is commenced. 

If the tailings are very acid, and a considerable proportion of 
the salts are found to be soluble in water, the general practice is 
to apply a preliminary water-wash. In order to neutralize the 
remaining acid, a quantity of the solution of caustic soda, equal 
to about 1 lb. of the salt to every ton of ore, is allowed to stand in 
contact with the tailings for two hours, and then drained off into 
the alkali-sump. 

The necessary quantity of caustic soda can be determined 
experimentally by the method recommended by Feldtmann. (See 
Chapter IV.) 

At the Witwatersrand Goldfields powdered caustic-lime is gener- 
ally used in place of soda. With very acid tailings as much as 
2£ lbs. to the ton is used. It is applied by adding the requisite 
amount to each truck-load before being dumped into the leaching 


The author generally used lime at the Government Experi- 
mental Works at the Thames, and found it preferable to caustic 
soda, as it is not attended with the production of ferrocyanide of 
zinc, which fouls the zinc in the extractor. On the other hand, it 
must be remembered that the addition of lime only improves the 
extraction when sulphides are not present in the material being 

When the cyanide contains sulphide, the extraction is certain to 
be slow and unsatisfactory. The sprinkling of lead acetate on the 
top of the tank will precipitate the sulphur of the sulphide as 
lead sulphide, thus leaving the cyanide free from sulphide. 

In the case of very acid tailings, Feldtmann strongly condemns 
the practice of conducting the preliminary washing in the leach- 
ing vat, on account of the possibility of the acid acting on the 
residual cyanide in the vat, and thereby liberating sufficient 
hydrocyanic acid to dissolve an appreciable quantity of gold, 
which he says would be lost, as it is not precipitated by zinc. 

He suggests that the washing should be effected in one vat, and 
the leaching in another; and considers that the extra costs of 
handling would be more than made up by the higher extraction. 

It is doubtful if hydrocyanic acid has any action on gold ; but 
it is probable that the ascending acid by uniting with any alkali 
present would form an alkaline cyanide which would at once react 
on any gold to which it had access. 


Strong Solution Leaching. — With dry-crushed ore, the 
strong solution, about one-third the weight of the charge, and 
generally containing from 0*2 per cent, to 5 per cent, of avail- 
able potassium cyanide, is allowed to pass into the vat from below. 
When the solution stands two inches above the surface of the ore 
the stop-cock is shut, and the solution lying below the filter is 
then drawn off, and nitration commenced. The strong solution 
generally takes from twenty-four to thirty-six hours to percolate 

In the case of tailings, the strong solution is added on top. 
After standing twelve hours, to allow the solution to permeate 
all the lumps of slimy material, the solution is slowly drained off, 
and passed through the zinc extractor. Small quantities are occa- 
sionally drawn off to assist diffusion of the solution through the 

It is a very common practice to apply a weak solution from the 
sump containing, say, 0*05 to 0*10 before applying the strong solu- 
tion, the intention being to allow the acids and foreign matters 


destructive to cyanide to expend themselves on the weak solution, 
before applying the strong solution. 

When the ore contains a large proportion of silver — say, from 
five to eight parts to every part of gold — it will be found neces- 
sary to adopt one of two courses in order to obtain a satisfactory 

Either a much greater bulk of potassium cyanide solution must 
be used to leach the ore — say, a quantity equal in weight to 
that of the ore taken — or else a much stronger solution must be 

The former, only, of these two courses would be applicable if 
the ore contained even a small proportion of copper oxide, car- 
bonate, or sulphide, or of sulphide of antimony, since the solu- 
bility of these in all solutions of cyanide, but especially in the 
strong, would render a fair extraction impossible, besides causing 
a heavy loss of cyanogen. 

The sulphide ores of silver are more slowly soluble in cyanide 
than gold, and for this reason the treatment of such is always 
more expensive than that of ordinary tailings or gold-bearing 
ores. On the other hand, the chloride of silver (kerate or horn 
silver) is more readily acted on than gold, the extraction generally 
exceeding 80 per cent. 

In the practical treatment of ores and tailings by cyanide, one 
of the first anomalies to attract the notice of the metallurgist is 
the fact that the strong solution, while it * loosens ' the gold, so to 
speak, yet does not carry it away, this being effected by the first 
and second cyanide washes. 

It is found that the first portions of the strong solution, draining 
from the charge, contain only from 0*02 per cent, to 0*1 per cent, 
of cyanide ; but the remaining portions come off stronger and 
stronger after a lapse of eight or twelve hours, until, towards the 
end of the strong leaching, the maximum strength is reached, 
after which the strength declines a little before the application of 
the first cyanide wash. 

The first portions of the solution during the strong cyanide 
leaching, being weak in cyanide and low in gold, are, therefore, 
passed through the weak zinc extractor, while the later portions, 
together with the first and second cyanide washes, are conducted 
to the strong extractor. 

It will be found in practice that fresh solutions are more active 
in dissolving gold than used solutions which in passing through 
the zinc-extractors become charged with the inert compound zinc- 

Strong Sump-Solution Wash.— After the strong solution 
has completely drained off, the strong sump-solution is applied 


from above, being run in on the top, or surface, of the ore. Its 
strength varies from 0*1 per cent, to 0*25 per cent, of KCy, and 
the quantity applied from one-third to one-quarter of the weight 
of the ore. 

This weak solution is sometimes allowed to slowly percolate 
through the charge, but more often it is drawn off as rapidly 
as possible, for the more rapidly the wash solutions are drawn 
off the more effective is the washing. The nitration is assisted 
by opening the stop-cock connecting the bottom of the leaching 
vat with the vacuum-cylinder. 

The percolation of the weak solution generally takes from twelve 
to twenty hours, the time depending on the condition of the pulp 
forming the charge. 

Weak Cyanide and Water Washes. — The number and 

strength of these will depend entirely on the character of the ore, 
or tailings, being operated on. In some cases it is found necessary 
to apply as many as three or four weak cyanide washes, from the 
weak sump, and then a final water wash ; in other cases the whole 
treatment consists of the strong cyanide leaching followed by a 
cyanide wash from the strong sump, a cyanide wash from the weak 
sump, and one or two final water washes. 

The quantity of each wash is in most cases one-half that of 
strong cyanide solution. 

The effect of the different washes should be carefully deter- 
mined by assaying the residues after each washing, and also the 
wash solutions as they drain from the vat. By this means the 
necessary number of washings will soon be ascertained. 

The assay sample should represent a fair average of the charge 
in the vat, and is easily and reliably obtained by taking a large 
number of cores, the full depth of the charge, by means of a tube 
shaped something like a cheese sampler. The scores should be 
dried and then sampled down for assay. 

Briefly summarized, the aim of the different solutions which 
have been applied to the material in the leaching vat is as 
follows : — 

(a.) Alkaline wash, to neutralize acidity so as to prevent 

waste of cyanide. 
(6.) Strong solution, to effect solution of gold, 
(c.) Weak solutions and water washes, to displace cyanide 
solution and to prevent cyanide from being thrown away 
with the residues. 
The different operations to be undertaken in the dry-crushing 
and direct method of treatment may be summarized as follows : — 

1. Dry crushing. 

2. Cyanide treatment. 


3. Copper-plate amalgamation of tailings. 

4. Concentration of tailings from plates. 

5. Treatment of concentrates. 

With wet-crushing the different operations are : — 

1. Wet-crushing. 

2. Copper-plate amalgamation. 

3. Concentration. 

4. Treatment of concentrates, by cyanide or other means. 

5. Treatment of tailings by cyanide. 

6. Treatment of slimes by cyanide. 

Remark. — When leaching slimy sands the successive solutions 
and washes should only be partially drained, just sufficient to allow 
assay samples to be taken. If very fine sands, or sands contain- 
ing slimes, be thoroughly drained, they become almost impervious 
to subsequent solutions. Hence it is important to partially drain 
so as to leave a cushion or jacket of solution round each particle 
of sand, the said jacket preventing packing and thereby permitting 
its easy replacement by the solution following. 



With all methods of wet-crushing and pulverizing the formation 
of a certain proportion of slimes seems inevitable, and when the 
ore contains metallic oxides, clay, or other earthy matter, the 
production is often very large. In ores containing a certain 
proportion of very fine or * float ' gold, the slimes are generally 
a valuable product, and their successful treatment has engaged 
the attention of metallurgists for many years. 

Since the introduction of the cyanide process, many attempts 
have been made to leach the slimes resulting from wet-crushing, 
on an economic scale, and the subject is one to which the author 
has devoted much thought and experiment for a number of years. 
The problem is principally a mechanical one, consisting in the diffi- 
culty of separating the solutions from the slimy mass, rapidly, 
effectively, and at such a cost as to permit the treatment of slimes 
of low value. 

Up to the present a great many different devices have been 
tried, with varying degrees of success. Among these may be men- 
tioned compression by hydraulic and other presses ; agitation and 
centrifugal force ; agitation and decantation ; and agitation with 
filtration, aided by an artificial vacuum. 

The dry-crushing of ores by means of the Californian stamp 
battery is always attended with the production of so large a 
proportion of slimes that only a shallow depth of the pulverized 
material can be leached in the direct method of cyanide 

The stamp battery was invented for wet-crushing, for which it 
stands unrivalled for all classes of ore. For dry-crushing it is 
a most unscientific machine, on account of its inability to dis- 
charge the dust when reduced to the requisite fineness. Never- 
theless, no better has yet been invented. 

In New Zealand and Australia every variety of pan and ball- 
mill, all the best known roll machines, and pulverizers of many 



different designs have been tried, but, except where the ore was very 
soft and friable, and the object is to produce slime, all have been 
j discarded for the stamp, which is the only machine known at present 

which is able to cope successfully with the hard quartzose ores. 

The depth of dry pulverized material placed in the leaching 
vats seldom exceeds three feet, even with the most favourable 
silicious ores. This necessitates a large plant to treat a compara- 
tively small output of ore, and a correspondingly greater cost per 
ton for treatment. 

Author's Experiments : — Early in 1893 the author treated a large 
parcel of ore from the Monowai mine, in the Thames district. 
The ore consisted of hard bluish and grey-coloured splintery 
quartz, containing a considerable proportion of sulphides of iron, 
copper, lead, and zinc. 

The ore was dried, dry-crushed, sampled and assayed, showing 
a value of £5, 5s. per ton. The crushing was effected in a stamp 
battery, which produced a large quantity of the finest slimes. 
These slimes rendered it impossible to effect the leaching by 
percolation, even with the aid of a vacuum. When mixed with 
water, the slimes, when only 4 in. thick, settled on the filter-cloth, 
I forming an impervious bed, through which it was found im- 

possible to draw off the solution. 

The pulp was then subjected to agitation, by which the dissolu- 
tion of bullion was effected in six or seven hours. The separation 
of the solution from the pulp, however, was a long and tedious 
operation, and extended over eight days. It was effected, but not 
very satisfactorily, by agitating the ore, allowing the slimes to 
settle, and then drawing the clear solution off by a syphon. The 
weak solution and wash- waters were added in succession, and the 
same operation performed after each. 

In order to ascertain the degree of fineness to which the ore 
was reduced when crushed, a number of experiments were made 
with a 60-mesh, 40-mesh, and an ordinary battery-punched screen, 
and it was found that the results were, in each case, as follows : — 

With 60-Mesh — 
A. 18% remained on ... 

C. 100% passed through a 

With 40-Mesh — 

A. 22% remained on ... . 

"' ™/o » » .... 

L». o/ „ „ .... 

D. 100% passed through a 

















With Punohed-Sorebn, equal to 30-Mesh- 

A. 26% remained on 

. 8100 sieve. 

B. 18% „• „ 

. 3600 „ 

C. 13^ „ ,, 

. 1600 „ 

D 4°/ 

900 „ 

E. 100% passed through a 

625 „ 

Taking the values of the different products separately, it was 
found that the finest in all cases gave the highest values. This 
also received confirmation from the circumstance that the fine 
dust, which had collected on an elevated platform during 
crushing, assayed higher than the dry material in the bin. 

The relative values were as follows : — 

From dust-bin, . . . .£550 per ton. 

„ platform, . . . 6 13 6 „ 

The ratio of gold to silver in the pulp from the dust-bin was 
nearly 1 to 9, and in the dust from platform 1 to 12, thus showing 
an increase of the silver contents. 

In 1893 the author obtained, with two of his assistants, a 
patent for a combined agitation and leaching process, which may 
be described as follows : — 

The appliances used in the operation consist of a shallow 
circular vat, a vacuum-cylinder, and an air-pump. The vat is 
provided with four revolving arms, to which soft rubber brushes 
are attached. The bottom of the vat is provided with a false 
bottom, consisting of a wooden grating covered with wool-packing 
or other webbing. The operation is conducted as follows : — The 
leaching solution, made up to the required strength, is first con- 
ducted into the vat. The revolving arms are then set in motion, 
and the dry pulp or fine slimes introduced. The agitation is con- 
tinued for six hours, or until the extraction is complete. A stop- 
cock, in a pipe connecting the false-bottom of the vat and the 
vacuum-cylinder, is then opened, and the air-pump started. The 
effect is immediate. The clear solution at once begins to drain 
over into the vacuum-cylinder, the brushes on the revolving arms 
preventing the slimes from settling and choking up the filter-cloth. 
When the slimes have been drained down to a thick paste, the 
first wash is added, the pump again started, and the slimes drained 
as before. The subsequent washes are applied in the same manner, 
and, when the washing is completed, a plug or door is opened, and 
the leached slimes are sluiced out. The whole operation of 
leaching and washing takes from eighteen to twenty-four hours. 

This process was adopted by the author for the treatment of 


several tons of ore from the Monowai mine, at Waiomo, with 
complete success. This ore contained a large proportion of clay 
and iron oxides, and, when dry-pulverized, formed a pulp which 
defied all the ordinary methods of percolation. Trial tests were 
also made with parcels of very fine slimes, and in all cases the 
results were most satisfactory. 

The author continued his experiments for the treatment of 
slimes in another direction, and in 1895, with Mr. G. W. Horn, 
obtained a patent for an improved leaching process for the treat- 
ment of slimes and other fine matters by cyanide or other solvents. 

The necessary apparatus consists of a solution vat, an air-com- 
pressor, and a leaching vat provided with a filter-frame and webbing, 
both at the top and bottom. The vat may also be provided with 
an air-tight cover, so as to permit the creation of an artificial 
vacuum to facilitate the filtration. 

In practice, the leaching vat, which may be constructed of wood, 
or any suitable material, and of any convenient size, is charged 
with cyanide solution. The slimes are then introduced and well 
mixed with the solution, which is then allowed to penetrate and 
stand in contact with them for two hours. At the end of that 
time the top filter-frame and webbing is put on, and additional 
cyanide, or other solution, is allowed to penetrate and permeate 
the slimes from below, thus displacing the solution already in the 
vat, and causing it to pass through the top filter-web as a clear 
liquid, which is conducted away by a launder. 

In order to prevent the solution selecting channels through the 
slimes, or other matter in the vat, during the upward filtration, 
the material is agitated by the discharge of compressed air through 
it from distributors placed in the sides and bottom of the vat. 

The gold in the slimes, being necessarily very fine, is dissolved 
very quickly, and the percolation can generally be started two 
hours after charging the leaching vat. When a portion of the 
gold is coarser than usual, or when it exists in the form of amal- 
gam, it is found that the agitation caused by the discharge of 
compressed air during the leaching greatly accelerates the 

The novel features of the process consist of upward filtration, 
and agitation by means of discharges of compressed air through 
the mass. 

The results of a number of small tests, on the finest slimes 
obtainable, have proved most satisfactory, and more extensive 
trials will soon be undertaken. One air-compressor would be 
sufficient for a number of leaching vats ; and the author is con- 
fident that the process is capable of very wide and useful applica- 



Acting on a suggestion thrown out by Mr. C. Wichmann, the 
author obtained from Mr. E. G. Banks, metallurgist to the Waihi 
Gold Mining Company, the results of a number of experiments, 
showing the relative values of the different degrees of fineness of 
the pulverized ore. These results are very instructive. 

Experiments showing degree of fineness and relative values to 
which "Martha" ore was reduced by stamping through 30- 
mesh Screens. 



0-36% remained on 30-mesh (900 holes) 
2-16% „ „ 40-mesh (1600 holes) 
9*29% „ „ 60-mesh (3600 holes) 

25-72% „ „ 80-mesh (6400 holes) 

74-28% passed 

Value per Ton. 
£4 2 6 

Similar Experiments with 40-mesh Screens. 

0*3% remained on 40-mesh (1600 holes) 
7-8% „ „ 60-mesh (3600 holes) . 

14-7% „ „ 80-mesh (6400 holes) 

85'3% passed „ „ 


18 10 
11 2 



3 12 7 
3 7 11 
3 11 4 
3 19 2 

Experiments showing value of Dust rising from stamp — 

Dry -crushing. 

Dust obtained from floor of mill to 10 ft. high 

10 ft. to 20 ft. high 
20 ft. to 30 ft. 
30 ft. to 40 ft. 
40 ft. to 60 ft. 









Value per Ton. 

£6 16 4 

7 2 4 

7 3 1 

7 19 11 

8 13 8 
4 1 4 



Average value of ore from which this dust came 

The different processes of slime treatment at present in use may 
be classified as follows : — 

1. Decantatiotl, as practised in Africa, America and New 

2. Filter-press method, as practised in Western Australia 
and New Zealand. 

3. Electro-chemical precipitation. 

4. Sun-drying. 

Decantation.— In 1896 Mr. C. Butters, then manager of the 
Rand Central Ore Reduction Company, erected a large slime plant 
at the Robinson mine, at the cost of nearly <£ 6 0,000. The salient 
features of the treatment are the classification of the slimes into 


three products. The first and second products, which consist of 
fine sands, are treated by ordinary leaching in vats; while the 
fine slimes are leached in vats with revolvi ug agitators, by a pro- 
cess patented in New Zealand by the author early in 1893. The 
agitators are about 10 ft. in diameter, and provided with a 
filter-frame nnd webbing. During agitation the solution is drawn 
off by means of an air-pump connected with a vacuum-cylinder. 
A fuller description of this process is given further on. 

On the Rand, the treatment of slimes has naturally been a very 
serious problem, to the solution of which much time and money 
have been devoted. The following method of treating slimes, 
known as the " natural settlement " process, has been in use at 
the Crown Reef since August 1896, with satisfactory results. It 
was first devised by Mr. J. R. Williams, and afterwards adopted 
by the Robinson Slime Works.* To the water, carrying the 
slimes from the separator plant, is added milk of lime, the slimes 
being thus precipitated in a flocculent form. The supply of lime 
is regulated by an automatic feeder, as too much lime is as bad 
as too little, since it interferes with the efficient precipitation of 
the gold. 

The slimes are settled in three large pointed boxes, two of them 
20 ft. x 20 ft. and 10 ft. deep, and the third 40 ft. x 40 ft. and 10 ft. 
deep. The settled slimes are drawn off at the bottom and pumped 
into the first two of eight treatment vats, about 90 per cent, of the 
water contained in them having been separated. 

The vats are each 32 ft. in diameter and 10 ft. deep, having a 
conical bottom. More water having been separated from the 
slimes by allowing them to settle, they are sluiced into a pump 
by a jet of cyanide solution and transferred to a second series of 
two vats filled with solution, and the strength of which is brought 
up to 0*1 percent, of KCy. 

About 80 per cent, of the gold is dissolved in the passage 
through the pumps, but agitation is continued from one to two 
hours by withdrawing the solution at the bottom and discharging 
it in oblique jets at the top and through the sides. 

The slimes are then allowed to settle, and the clear solution is 
drawn off through side-cocks (which have been replaced by syphon 
pipes at the Robinson Works) and passed to the precipitation boxes. 

The residual slimes are then pumped successively into the 
third and fourth series of two vats, where they are further 
agitated with very dilute solutions of cyanide and allowed to 
settle. These solutions do not pass to the precipitation boxes, 
but are transferred to the preceding series of vats. 

* Chem. and Met. Soc. of S. Africa, July 1897. 


The "strong" solution from the second series of vats is run 
into two settling tanks, 15 ft. in diameter and 5 ft. deep, where 
it is allowed to clarify. Electrical precipitation is used. From 
6643 tons treated, an actual extraction of 60 '5 per cent, was 
obtained, at a cost of 3s. 9d. per ton. 

The concentration of the solution as regards its gold contents 
when decanted, was first successfully introduced by Mr. J. R. 
Williams at the Crown Reef plant, in order to reduce the amount 
of solution passing through the precipitating boxes. Naturally, 
when a given amount of gold in solution is obtained from the 
slimes per twenty-four hours, if this amount of gold can be con- 
centrated into a small volume of solution, a small precipitating 
capacity is required. For these reasons Mr. Williams introduced 
the system of double washing ; that is to say, the first solution 
that is applied to the slimes after decantation is not clarified and 
passed through the precipitation boxes, but is run into what is 
called an " intermediate tank," and then pumped up for use again 
upon a fresh charge of slimes. After this second decantation, it 
is run through the clarifying apparatus to the boxes for precipita- 
tion. The settled slimes are then re-pulped with an equal volume 
of solution from the precipitated sump ; that is, solution which 
has passed through the precipitation boxes. Theoretically, the 
enrichment of the unprecipitated solution could be carried on 
until it was equal to the value of the slimes, but in practice it has 
been found that the first wash which has not been precipitated, 
and the second wash which has been precipitated (two washes 
altogether), make the simplest and most perfect system. 

South African Practice. — The main features of the decanta- 
tion process as practised in South Africa are as follows : — 

1. Spitzlute separation of sands and slimes. 

2. Spitzkasten concentration of slimes. 

3. Collection and settlement of slimes in collecting vats pro- 

vided with decanting syphon. 

4. Agitation of slimes with cyanide solution. 

5. Settlement of slimes and decantation of solution for precipi- 


At the Rand mines, the agitation is generally effected by a 
centrifugal pump, but in some cases revolving stirrers are used. 

New Zealand Practice. — The method of slimes treatment 
adopted by the Waihi Company at their new mill at Waikino in 
1901, is based on the best South African practice, differing only in 
the separation of the gold-containing solutions which are separated 
from the slimes by filter-presses instead of repeated decantation as 
at the Rand. 

A detailed description of the methods of slime treatment at the 


Waihi Company's mills of 320 stamps, furnished by the company's 
metallurgist, Mr. E. G. Banks, will be found in Chapter XII. 

Filter-Press Practice. — This process has been revived in 
New Zealand during the past few years, but it is in Western 
Australia that it has received its widest and most successful 
application. The scarcity of water and the clayey character of 
the oxidized ores of that tropical country presented a difficult 
problem to the metallurgist and chemist. Compared with the 
decantation process, a higher average extraction is being obtained, 
with a lower consumption of water ; but the assay value of the 
residues is higher than in South Africa, where it has been re- 
duced as low as 9 J grains of gold per ton. A consideration of the 
practice shows that the higher extractions at Kalgoorlie are due 
to the higher grade of the slimes treated. 

Details of Process. — The following details of the process as 
practised in Western Australia are extracted from an instructive 
paper recently published by Mr. Donald Clark, Director of the 
Bavinsdale Mining School in Victoria.* 

The ore at Kalgoorlie, as a rule, is dry crushed, Krupp's ball 
mills being mainly used ; after crushing, it is wholly roasted. 
Several types of furnaces are on the field, including the shaft 
furnace (similar to those at Mount Morgan), Brown's, Ropp, 
Holthoff-Wethey, and Edwards' mechanical furnace. It is note- 
worthy that the last is most valued, on account of its lowness in 
cost and its perfection in roasting. 

The analysis of typical ore from the Lake View Consols mine is 
appended : — 

Before roasting. 
Insoluble . 62*80 

After roasi 


FeS . . ** 8-27 

Fe 2 3 


CaC0 8 . . 13-89 

CaO . 


MgCO, . 7 37 

MgO . 

SO, . 


S (undecomposed) '11 

These results show that the ore is not a complex one, since it con- 
sists of the usual gangue and a comparatively small percentage of 
pyrites. The presence of telluride of gold gave rise to the 
greatest difficulties in the treatment of this ore. 

The method finding most favour is to pass the whole of the 
roasted ore into a hydraulic separator, the fine gold and slimes 
flowing over the top, while the coarse gold, the partly roasted 
sulphides, and the coarse sands are drawn away at the bottom. 
The slimes are led into spitzkasten, where 50 per cent, of the 

* Clark, The Australian Mining Standard, Dec. 5, 1901. 


water flows off at the top clear, and the mud, containing 50 per 
cent, of dried slimes, is run into agitators, where it is agitated 
with cyanide solutions and afterwards filter-pressed. The coarse 
sands from the hydraulic separator are run over copper plates, 
then over percussion tables, where the coarser concentrates are 
extracted. The sand which escapes from these is run into a vat, 
and there treated in the usual way with cyanide solutions. The 
concentrates from the table are ground to slimes in an amalga- 
mating pan ; the slimes are sent to the agitators, thence to the 

The Dehne filter-press is the favourite press in use and has 
given great satisfaction. The Martin press has lately been placed 
in the market and is worked as follows : — 

The slimes are forced in through the slime valve and passage, 
and thence find their way through the ports or side openings of 
the open frames which they fill up, the liquor escaping by the 
drain cock. 

When the water ceases to run, these plates are full of fairly 
caked slime The cocks are then closed. Solutions for cyaniding 
or washing, as the case may be, are then forced in through the 
solution channels. These find their way into the solution plate, 
and are forced from the corrugated plate through the perforated 
one, then through the slime cake, through the filter-cloth, the 
perforated plate, and down the corrugated surface of air-plate 
whence they are led away to the zinc boxes. It should be stated 
that any imprisoned air is first got rid of by opening the air escape 
cocks, and displacing the air by the solution. When the liquid 
starts to run through the air-cock, the solution plate is closed. 

When cyaniding in presses, the escaping water is clear enough 
to be led back to the separators ; the cyanide solutions are forced 
through at a pressure of 90 lb. per square inch until the gold is 
extracted. It takes 90 minutes to fill, 90 minutes to extract the 
gold, and 90 minutes to empty three such presses. In order to 
displace any wash water or solution, a current of compressed air is 
turned on to the solution passage for a few minutes, and the press 
is ready for opening. The cakes or slabs of slime are usually 
3 ft. square, and a 50-cake press 2 ins. thick contains 80 cubic feet, 
or about three tons. Messrs. Martin & Co. are now making 
presses 42 in. square for 3-inch cakes, the 50-cake press holding 
nearly six tons. When the presses are used in conjunction with 
the agitators, as many as from 10 to 12 charges may be put 
through in twenty-four hours, or, in other words, as much as 60 
tons per day. According to Mr. J. Moss, the cost for filter 
pressing without any grinding power is 10s. 10'2d. per ton, the 
average residues being 1 -95 dwt. per ton. The total cost of treat- 


ment from the mine to the tailings dump, allowing for first cost 
in labour and supplies and second cost in repairs and renewals, is 
put down at 35s. 9'4d. per ton. 

The Dehne machines can, if necessary, be constructed of gun- 
metal or wood, with all the channels and valves lined with 
material suitable for withstanding the corrosive effects of any 
particular material. 

Details of the Diehl and Riecken slime processes, both of which 
are in operation at Kalgoorlie, will be found in Chapter XIII. 

Roasting previous to Cyaniding. — With pyritic concen- 
trates and even high grade pyritic ores, a higher percentage of the 
gold contents can always be obtained by roasting previous to 
cyaniding, and in the Cripple Creek camp of Colorado and 
Kalgoorlie in Western Australia this process is largely adopted 
for the treatment of sulpho-telluride ores. 

By roasting, the cyanide solutions are kept freer from soluble 
salts than when treating raw ore ; a higher percentage of gold can 
be extracted at a smaller cost ; the time of treatment is shorter ; 
and it is found that in the case of clayey ores, roasting causes 
dehydration, thereby rendering them porous, and making filtration 
comparatively easy. 

The author found that Moanataiari concentrates, which yielded 
Only 30 per cent, of their values in the raw state, yielded, when 
roasted, 90 per cent., with a smaller consumption of cyanide. 

In order to ensure success it is essential that the sulphides 
should be subjected to a "dead" roast. 

A quick and satisfactory test to determine if the ore is dead 
roasted, and amenable to cyanide treatment, is described by 
Wallace Macgregor as follows * : — 

" Take samples of the roasted ore at discharge end of furnace, 
cool, take from 100 to 250 grams, place in a beaker with 200 c.c. 
of water ; stir this by shaking for about a minute, then filter into 
a beaker or flask, and to the filtrate in the beaker add a small 
quantity of cyanide solution made up to the same strength as that 
used in the regular work of the plant. It is best to add the 
cyanide solution slowly and carefully, noting the result. If no 
cloudiness at all appears the ore is dead roasted, or at least well 
fitted for treatment by cyanide solution, and the consumption of 
cyanide will be normal. 

" If a brown coloration takes place there are still some soluble 
salts of iron left in the ore, which will cause a somewhat higher 
consumption of cyanide, and may lead to a precipitation of com- 
pounds of ferro-cyanide in the zinc boxes. If, on addition of 

* Engineering and Mining Jowrnal, and Mineral Industry , 1898. 


cyanide to the filtrate, a blue coloration, soon becoming a 
greenish blue precipitate, is formed, then the ore is very badly 
roasted, and one may look for a high consumption of oyanide, and 
the circulating solution will be made foul." 

As to the economy of roasting ores, that must be determined 
for each individual case. There are oxidized ores and tailings 
from stamp mills which pay a fair profit by direct treatment with 
cyanide. These ores and tailings may be so low-grade that roast- 
ing would be out of the question, although if used it would increase 
the extraction of the gold on subsequent treatment with cyanide 
by probably 10 per cent, or more. There are cases of heavy sul- 
phuretted ores and concentrates where roasting interferes with the 
extraction of the gold, and causes a very heavy consumption of 

The advisability of roasting ores previous to cyaniding should 
be carefully determined in each case, both by laboratory experi- 
ments and small plant tests, where 1 to 10 tons of dead roasted 
ore could be worked. 

The main features of the cyanide process used for the treatment 
of the sulpho-telluride ores consist in drying the ore ; dry-crush- 
ing in stamps, Griffin or other mills ; roasting the pulverized ore 
in furnaces; amalgamation in pans; cyaniding sands, if any, in 
vats by percolation and slimes in agitators ; filter-press separation 
of solutions from slimes, and final washing of slimes. 

A detailed description of the plant and process used at some of 
the leading Kalgoorlie mines will be found in Chapter IX. 

In the Ohinemuri Goldfields of New Zealand, the gold occurs 
in an extremely fine state, and hitherto dry-crushing and 
cyaniding have been used, the objection to wet-crushing being the 
difficulty experienced in treating the slimes which were always 
high grade. 

The successful experiments conducted at the Crown mines with 
wet-crushing last year, and the subsequent adoption of this 
process at the Company's mill, have shown that, with careful 
manipulation, the simultaneous treatment of the slimes can be 
carried on with very little extra cost. 

The following description of wet-crushing with cyanide solution 
has a special interest.* The ore is delivered to the rock-breaker 
without drying. With the ore there is fed into the mortars a 
constant stream of a stock cyanide solution of about '1 per cent, 
strength, regulated by stop-cocks near the boxes. The rate of 
flow is kept between 35 cwt. and 2 tons per hour, being gauged 
at the supply-tanks. As the stamps, which weigh 1000 lbs. with 

* Thorpe, Auct. Min. Standard, Jan. 19, 1899. 


96 drops to the minute, crush about 2 tons per stamp per day, 
the amount of solution is consequently about equal to the ore 
crushed. Previous experiments showed that a larger proportion 
of solution, while it increased the output, caused the pulp as well 
as the gold to be washed through the screens in a state too coarse 
for cyanidation. From similar considerations the discharge is 
raised to a height of 8 or 9 inches, the screens being 40-mesh. 
From the front of the boxes the pulp flows to a wheel-and-bucket 
elevator, and thence to the vat-shed. The vats are of the size 
usual in dry-crushing on those goldfields, viz., 22£ ft. in diameter 
and about 4 ft. in depth, not including the space of 4 inches below 
the filter-cloth. The pulp flows into the centre of a wooden, two- 
armed distributor, driven by overhead gearing. Suspended to 
the arms are five sets of rectangular pieces of sacking, each set 
being composed of three or four pieces of varying lengths, so that 
they may be lifted as the vat fills. Their object is to keep the 
slimes on the surface of the sands from settling. These slimes 
are being continuously decanted off by an adjustable pipe into a 
receiving vat on the same level, provided with an agitator similar 
to that used in an ordinary settler. Thence they are pumped 
from time to time by a small centrifrugal pump into the settling 
vats, also provided with light agitators. These vats are similar 
to the sand vats, but have the filter-cloth and frame removed. 
From these the clear solution is kept continually draining into a 
sump, whence it is pumped to the head of the precipitating boxes. 
After passing through the boxes it is pumped into the supply 
tanks, whence it feeds the stamps or is used for washes. 

The sands, which contain a certain amount of slimes, are treated 
as ordinary tailings, percolating, with a vacuum of 20 inches, at 
the rate of about 5 to 8 tons per day. The extraction obtained 
is over 80 per cent., and with a few modifications in the treatment 
will probably be better still. The strength of the strong solution 
is '5 per cent., and the solution draining off is run through the 
towers with that coming from the slimes. In this way the stock 
solution is kept up to about *1 per cent, in strength, with little 
variation. The slimes give no difficulty in the solution of gold, 
the '1 per cent, solution being apparently quite strong enough. 
Each slime vat is filled with slimes till the latter settle down 
to about the middle of the vat, the proportion of solid matter 
in this state to the solution being about 13 : 20. The agitator 
is then started, and a weak wash added, equal, if possible, to the 
amount of slimes. In half an hour these are sufficiently mixed, 
the agitator is stopped, and the clear solution is gradually 
decanted off by adjustable, fan-mouthed pipes. In this way, 
neglecting, for the sake of simplicity, the amount of solid matter 


in the slimes, the original amount of gold solution in the vat 
is reduced to a half. A second wash reduces it to a quarter, 
a third to an eighth, and so on. In this particular instance about 
20 tons of slimes are given four washes, or from 15 to 20 tons 
of cyanide solution each, and one of water. The extraction of 
gold usually amounts to 90 per cent., a few exceptions being 
probably due to the imperfect decanting of the slimes from the 
sand vat, from which some of the coarser material might easily 
descend into the slime vat. 

The method of working is as follows : — The ten -head employed 
in this process fills one vat with sand in three days. There being 
three sand vats, and the stamps being " hung up " during Sunday, 
seven days are available for treating the sand in each vat. For 
treating the slimes there are four vats. This allows the slimes 
from a day and a half's crushing to be put into one vat, each 
of the four vats being filled in the six days. This gives five and 
a half days for washing, etc. The washes decant off, with the 
aid of lime, at the rate of from 15 to 20 tons per day. The 
extraction of gold in the precipitating boxes is so far very 
good. The chief advantages of this method over dry-crushing 
are: (1) Elimination of the expense of drying the ore; (2) 
increase in stamp duty ; (3) economy in consumption of cyanide ; 
(4) saving of labour in filling ; (5) absence of dust. In the mill 
under review the stamp duty has been increased from one and a 
half tons to two tons per day per stamp ; the consumption of 
cyanide decreased by about half. The "cyanides" formed in 
roasted ore are absent, while the strength of the solution used in 
the boxes is too weak to suffer much loss by agitation. Crude as 
are some of the appliances used in this process, the results justify 
the expectation that dry-crushing with cyanidation will soon be 
entirely displaced by wet-crushing with cyanide. 

Sun-drying Slime Process. — At the Try Fluke mine at 
Kuaotunu in New Zealand, the slimes were dried in the sun, 
broken up, and then mixed with sands in the proportion of one 
part of slimes to two parts of sand. This method gives very 
satisfactory results, but can only be carried on in a very dry 

This slime process is in use in India, and in the United States 
at the Dexter plant, at Tuscarora in Nevada, where one part of 
dry slime is mixed with one part of sand and leached with a 
0*20% solution of cyanide for three to five days, the solution 
running in at the top as fast as it runs out at the bottom until 
an adequate extraction has been obtained. 


With low-grade ores, it is generally more advantageous to omit 
close concentration and 'rather to classify prior to cyanide treat- 
ment. In all oases the best results are obtained from the cyanide 
treatment of concentrates when the pyrites and sand exist in the 
proportion of five to one. 

Pyritic concentrates may be leached by agitation with cyanide, 
or simply by percolation, as with ordinary tailings. Both methods 
give satisfactory results. 

Leaching by Percolation. — At the Watersrand Goldfields the 
treatment of concentrates by cyanide has been largely adopted in 
preference to chlorination. 

From the storage vats the concentrated material is taken to the 
leaching tanks, and subjected to the action of strong cyanide 
solution, containing from 0'4 per cent, to 0*6 per cent, of cyanide, 
for periods varying from twelve to eighteen days. In practice, 
the solution is allowed to slowly percolate through the concen- 
trates, and it is then passed through the zinc precipitation boxes. 
It is again made up to the original strength, and allowed to 
percolate as before. This operation is continued until a satis- 
factory extraction is obtained. At the Crown Reef Cyanide 
Works the cost of this method of treatment is said not to exceed 
17s. per ton of 2000 lbs. 

Mr. C. M. P. Wright gives the following useful details of the 
cyaniding of concentrates by percolation at the Choukpazat gold 

" The concentrates consist of 30 to 40 per cent, sulphides and 
60 to 70 per cent, coarse sands : of the sulphurets, other than 
iron pyrites, 5 per cent, consists of franklinite, galena, chal- 
copyrite, and a very little altaite. Of these minerals the franklin- 

* Wright, Inst. Min. and Met., London, Nov. 20, 1902. 



ite is by far the richest, assaying from 7 oz. per ton and upwards ; 
the galena holds practically no gold, and the chalcopyrite and 
iron pyrites vary from about 18 dwt. per ton to 2 oz. per ton, 
depending on the general value of the ore. 

" Our mode of treatment is simple. After an alkaline or plain 
water wash, follows a weak solution wash (catch -sump strength 
0*10 per cent, to 0*12 per cent. KCN) and then nine washes of 
0*3 per cent, strength : the contents of vat are carefully turned 
over, and 0*3 per cent, solutions follow until for two successive 
days the effluent comes down to 0'26 per cent., when the treat- 
ment is considered complete. Two more solutions, catch strength 
of strong solution sump, usually 0*25 per cent., are followed by 
two weak solution 0*07 per cent, washes, and by a final water 
wash, or two if required. 

" Every solution and wash that passes through the percolation 
vat passes into the same sump through the same zinc box. A 
careful check is kept of cyanide used, of solutions taken from 
weak solution sump, and of strong solution thrown on to tailings 
vat so that the level of the special sump is kept constant and at 
same time the somewhat loaded strong solution is exchanged for a 
cleaner weak solution. 

" The check upon consumption of cyanide and zinc is complete 
and automatic. Usual tonnage charged monthly, 17 to 20 tons ; 
duration of treatment, 24 days; extraction, 84 per cent. 
Formerly the zinc in zinc box was made up mid-monthly ; this 
was found to be not necessary and wasteful of zinc. This zinc 
box is now carefully packed at the commencement of the treat- 
ment and left absolutely untouched till the clean-up: it is an 
eight compartment 1,000-ton per month box, and each compart- 
ment is filled." 

Mr. Wright states that owing to the presence of copper in the 
concentrates, all the zinc became coated with that metal immedi- 
ately after being put into use. The cost per ton of concentrates 
treated was 12s. 10d., and the extraction 84 per cent. 

Leaching by Agitation. — At the Woodstock mine in New 
Zealand, the concentrates are treated by agitation in small 
wooden vats provided with mechanical stirrers fixed on a vertical 
steel shaft which is actuated by bevel-gearing, which, in turn, 
derives its motion from an intermediate shaft by means of a belt 
and suitable pulleys. 

The concentrates have a value of £30 to £40 per ton, a large 
proportion of the value being in silver sulphide (argentite). They 
are agitated for thirty-six hours with a 4 per cent, solution of 
cyanide. Two pounds of lime are added for every ton of concen- 
trates. The charge weighs 1£ tons. The actual recovery is said 


to vary from 90 to 94 per cent, at a cost of 18s. per ton for 
labour and material. 

The leaching of concentrates by agitation was first introduced 
in New Zealand at the Sylvia mine by Dr. Scheidel in 1891, 
where the results were very satisfactory. Details of the plant 
and treatment will be found in the next chapter. 



The first attempt to introduce the cyanide process on a working 
scale, for the recovery of gold and silver from their ores, was made 
by the Cassel Gold Extracting Company, of Glasgow, at the New 
Zealand Crown mines, Karangahake, in 1889. The operations 
were under the supervision of Mr. J. M c Connell. In the first 
plant, agitation formed a prominent feature, but in later years 
leaching by percolation became the more general and favourite 
method of treatment. 

The new and extensive cyanide works of the Crown Mines Com- 
pany, besides a percolation plant of twenty-four tanks, contained 
an agitation plant consisting of sixteen wooden tubs, or barrels, 
fitted with revolving paddles. The agitators were seldom used, 
preference being given to percolation. 

At the concentration plant at the Sylvia mine, Thames Gold- 
field, an agitation plant was erected by Dr. A. Scheidel, in 1891, 
consisting of three agitators, 6 ft. in diameter and 6 ft. deep; 
three vacuum -filters, together with the necessary solution tanks, 
zinc extractors, and other appliances for cyanide treatment. 

The ore was heavily charged with iron pyrites, and occasionally 
a small proportion of zinc-blende, argentiferous galena, and copper 
pyrites. It was wet-crushed in a 10-stamp Calif ornian battery, 
classified in pyramidal boxes, and subsequently concentrated in 
jiggers, slime-tables, and buddies. 

The concentrates of four grades were afterwards subjected to 
cyanide treatment by agitation. The extraction from the best 
slimes is said to have amounted to 96 '45 per cent, of the gold, and 
94*59 per cent, of the silver. The average extraction from all 
classes of concentrates amounted to 82*67 per cent, of the assay 

The strength of the cyanide solutions varied from 0*5 per cent, 
to 1 per cent., and the time of agitation from six to twenty-four 

* The Cyanide Process, by Dr. A. Scheidel, p. 79. 




hours. About 300 tons of concentrates were treated, and yielded 
over £10,000 value of bullion, but the cost per ton is not given. 

The extraction is said to have been satisfactory until an excess 
of copper ore appeared in the concentrates, which rendered them 
unsuitable for cyanide treatment. 

The cyanide plant at the Thames School of Mines, designed and 
erected by the author, is provided with an agitator which serves a 
double purpose, being used also as a dissolving vat when ores are 
being treated by percolation. 

The agitator is similar to those used at the Crown and Sylvia 
works, consisting of an upright tub or barrel provided with a 
central revolving spindle set in a foot-step at the bottom. At the 
bottom end of the spindle is fixed a screw, consisting of four 
paddles or blades. The foot-step, being in the agitator, is subject 
to great wear and tear, and this forms a most objectionable 
feature, as it must be continually renewed. This difficulty is 
easily got over by fixing a hollow cone in the centre of the 
agitator and placing the spindle in this, the motion being applied 
from below, as in most grinding and amalgamating pans. An 
agitator of this kind was erected in 1894 by Dr. Scheidel at the 
Utica mine, Calaveras County, in California.* 

The treatment of concentrates by agitation has already been 
described at the end of the preceding heading. 

The Actual Extraction by Agitation. — The practice of 

the author was to charge the agitator with the cyanide solution 
made up to the required strength, using from 40 per cent, to 
60 per cent, of the weight of the ore. The agitator was then set 
going, about fifty revolutions per minute, and the ore or tailings 
gradually fed in, until the charge was complete. The agitation 
was continued until a satisfactory extraction had been effected, 
which generally took from six to ten hours. Samples for assay 
were obtained from the agitator by means of a small tin at the 
end of a stout string. When the extraction was considered 
adequate, the agitator being still in motion, the stop-cock was 
opened and the pulp allowed to discharge into a percolation vat, 
where the solution was drawn off and the residues washed, aided 
by an artificial vacuum. The cyanide solutions were then passed 
through the zinc precipitation boxes. 

Leaching by agitation possesses many advantages over percola- 
tion for the treatment of pyritic concentrates or rich tailings. The 
dissolution of the gold is much more expeditious, taking hours 
where percolation requires days ; and with suitable material the 
extraction is always high. 

* The Cyanide Process, by Dr. A. Scheidel, 1894, p. 89. 




On the other hand, agitation requires motive-power, and from 
the nature of the process the charges must be small, in no cases 
exceeding a few tons. There is a prevalent belief that agitation 
causes an excessive consumption of cyanide by decomposition by 
atmospheric carbonic acid gas ; but the author thinks this source 
of loss much exaggerated, and is certainly much less than it was 
in the early attempts at cyanide treatment when agitation was 
prolonged from thirty-six to forty-eight hours continuously. 

The author has found, by many trials, that from six to eight 
hours' agitation is sufficient to effect the dissolution of the gold in 
the most refractory ores when reduced to a sufficient degree of 

Numerous experiments during the progress of agitation have 
shown that the greater portion of the gold was dissolved during 
the first hour. 

The rate of the extraction at the different hour-periods, during 
the treatment of the Monowai sulphide ore, is given below, and 
will be found instructive : — 

Time of Agitation. 

Gold Extraction %. 

After 1 hour 

85 4 

,, 2 „ 


„ 3 „ 


,, 4 „ 


„ 5 „ 


„ 6 „ 


From the above it will be seen that the maximum extraction 
was obtained in four hours. This was a complex sulphide ore, 
containing sulphides of copper, zinc, iron, and lead. An analysis 
of the more mineralized portion by Mr. F. B. Allen, M.A., B.Sc, 
Director of the Western Australia School of Mines, gave the follow- 
ing results : — 

Insoluble gangue, . . . .90*15 

Copper pyrites, 
Iron pyrites, 
Alumina, . 
Water and loss, 




The bullion contents of this ore were, gold 1 oz. 5 dwts., and 
silver 14 oz. per ton. A 0*6 per cent, solution of cyanide was used 
for leaching by agitation, and the consumption amounted to 0'45 
per cent., with an extraction of 92 per cent. A very large amount 
of copper was dissolved, and, becoming deposited on the zinc, 
caused much trouble in the precipitation of the bullion. The 
actual extraction was below 70 per cent. 

The following interesting and instructive experiments by agita- 
tion with 0'25 per cent, and 0*33 per cent, cyanide solutions on 
" Martha " ore were kindly supplied by Mr. E. G. Banks, the 
metallurgist for the Waihi Gold Mining Company at Waihi : — 

Experiments on " Martha " ore, showing rate at which the precious 
metals were extracted by a 0*25 'per cent, solution of KCy, and 
amount of KCy consumed. 


dwt. gr. 



per ct. 


Ore Before Treatment, 

16 8 

3 4 13 

1. , 

, after 2 hours' agitation, 

9 19 

2 15 20 




2. , 

, „ 4 


8 4 

2 13 22 




3. , 

„ „ 6 


6 14 

2 7 10 




4. , 

, ,, 8 

t > 

4 14 

2 4 40 





,, „ io 

1 1 

3 12 

1 19 20 





» ,. 12 


3 12 

2 3 1 





,, „ H 

> > 

3 6 

2 2 11 





» „ 16- 

; i 

3 6 

2 2 11 





„ „ 18 


3 10 

2 3 3 





., ,, 20 

j j 

3 1 

2 4 8 





„ „ 22 


2 15 

2 2 8 





„ „ 24 


2 11 

2 2 11 




Fineness of 

Ore Treated, 


remained on a 40-mesh screen. 








The zinc for bullion precipitation is used in thread-like turnings, 
as this form gives the most surface for the least weight. It 
should be free from arsenic or antimony, although a little lead is 
an advantage, as it causes more rapid precipitation by forming a 
voltaic pair with the zinc. 

As a general rule, one cubic foot of zinc turnings will precipi- 
tate the gold from two tons of solution in twenty-four hours. 

Zinc on which a film of bullion has been precipitated is more 
active than pure zinc, and it is therefore advisable to replace the 
zinc dissolved in the upper compartments by moving the zinc 
forward from the lower compartments, and adding fresh zinc to 
the latter. 

In practice, the cyanide solution is allowed to slowly drain 
through the zinc in the precipitating boxes. The rate of flow is 
soon determined by actual experience. It is generally found that 
85 per cent, to 95 per cent, of the bullion will be precipitated in 
the first three boxes. 

The solution, after leaving the boxes, should not contain more 
than six or eight grains of gold to the ton. 

Zinc, in the form of zinc-dust and zinc-fume, is in use for pre- 
cipitation, particularly the former, but neither seems likely to 
supersede zinc turnings, which possess the important advantage 
that they afford a continuous method of precipitation, whereas 
dust and fume have to be applied in charges in solutions collected 
in separate vats. Furthermore, the zinc extractor-boxes, when 
once in good working order, fequire very little attention except at 
the periodical " clean-up." 

The principle of precipitation of gold by metallic zinc is based 
on the fact that cyanide has a stronger affinity for zinc than for 
gold, as shown by the following equation : — 

2AuKCy 2 + Zn = ZnK 2 Cy 4 + 2 Au. 



By the above reaction it will be seen that 1 oz. of zinc should 
precipitate 6 oz. of gold, but in practice it is found that from 4 
oz. to 12 oz. of zinc are required to precipitate 1 oz. of gold. The 
reactions which take place in the zinc precipitating boxes are at 
times most varied and perplexing, especially during the treat- 
ment of pyritic tailings or acid mineralised ores. 

Part of the excessive consumption of zinc is no doubt due to 
decomposition by free cyanide, as may be ascertained- by testing 
the solution for available cyanide before entering and after leav- 
ing the zinc precipitating boxes ; but the consumption and conse- 
quent loss of cyanide by this cause is much less than generally 
supposed, and in all cases insufficient to account for the great 
waste of zinc. 

Zinc is soluble in an aqueous solution of potassium cyanide 
without evolution of hydrogen. In the extractors, the cyanide 
comes in contact with so extensive a surface of zinc that a large 
quantity of that metal must pass into solution, but curiously 
enough the consumption of free cyanide in the extractor does not 
correspond with the consumption of zinc, and we can only con- 
clude that a process of regeneration takes place in the extractor. 
It is quite certain that the fouling of cyanide solutions with that 
troublesome inert alkaline substance, zinc-potassium-cyanide, takes 
place iii the passage of the solutions through the extractors. 

It might naturally be expected that zinc would accumulate in 
the cyanide solutions to a detrimental extent, but this is found in 
practice not to be the case. The zinc does not accumulate to 
any extent, a result in all probability due to the action of the 
sulphides contained in the ore and cyanide which cause its 
precipitation as a sulphide of zinc. 

The precipitation of the gold, doubtless from electro-chemical 
causes, is always more rapid and complete from moderately strong 
than from very weak cyanide solutions, but under all circum- 
stances the solutions must be distinctly alkaline to ensure a satis- 
factory precipitation. 

It has been suggested by some chemists that this is due to 
nascent hydrogen, liberated by the action of the free KCy on the 
zinc, taking the place of the gold, according to the following 
equations : — 

4 KCy + Zn + 2H 2 = ZnK 2 Cy 4 + 2KHO + H 2 

2 AuKCy 2 + H 2 = 2KCy + 2HCy + Au 2 . 

The liberated hydrocyanic acid is capable of combining with 
any free alkali present, and thus there would be no loss of the 


cyanogen combined with the gold. This reaction is shown by the 
following equation : — 

HCy + NaHO = NaCy + H 2 0. 

Hydrogen gas is always evolved when gold is precipitated, 
and the gentle action of the gas bubbles, as they rise to the 
surface in the zinc boxes, is an indication of satisfactory pre- 

During the treatment of pyritio tailings at Kuaotunu, the un- 
satisfactory precipitation of the gold was for some time a source 
of much trouble to the chemists in charge of the cyanide works, 
but this difficulty was overcome by making up the strength of the 
solution before entering the extractor to something like the 
original working strength. 

In practice, this was effected without any extra trouble by 
simply placing a barrel containing a strong solution of cyanide at 
the head of the extractor, and allowing a steady drip into the 
cyanide solution, in the top compartment, which was filled with a 
filter of sand and gravel. By testing the cyanide solution a few 
times, the rate of drip to bring it up to the required strength was 
easily determined. 

The author used this method with success in the treatment of 
cupriferous ores from the Monowai mine in the Hauraki Gold- 
fields, in 1895. 

It was found that the dissolved copper was precipitated much 
more rapidly from a weak solution of cyanide than from a strong 
one. In order to overcome this, the solutions were made up to 
the original working strength. This method has now been super- 
seded by the lead acetate pickling process. 

In the treatment of slimes by decantation and in filter-press 
processes, but especially in the former, there are formed large 
volumes of excessively weak cyanide solutions containing gold. 
It was a matter of early cyanide experience that zinc precipitation 
from such weak solutions was very imperfect, in fact far inferior 
to electrical precipitation. 

The discovery was soon made that the lead-couple used for the 
precipitation of gold from solutions containing copper was also very 
effective for the precipitation of gold from extremely dilute 
cyanide solutions, and the practice in South Africa and elsewhere 
in slimes plants is to pickle the zinc, before use, in a trough 
containing a 10 per cent, solution of lead acetate, until all the 
zinc is covered with a black coating. The precipitation effected 
by this lead-couple is almost perfect, only a trace of gold, as a 
rule, finding its way into the sump. 

At the Camp Bird mines, as described in the following page, a 


zinc-mercury couple is used with very satisfactory results for 
solutions containing copper, and is stated to give better results 
than the lead-couple. 

It is the practice at some plants to allow a dilute solution of 
lead acetate to drip slowly into the head of the zinc extractors, 
but the practice is not to be commended, on account of the 
difficulty of regulating the rate of flow. Besides, by this method 
an excessive amount of lead is certain to find its way into the 

It is now the practice when the cyanide contains sulphides to 
sprinkle lead acetate on the top of the tank. By this means the 
alkaline sulphide is decomposed, and the cyanide freed from 
sulphide by the precipitation of lead sulphide. 

When copper is present in the solution, it soon covers the zinc 
with a bright metallic coating, which begins in the lower boxes, 
and gradually encroaches on the upper ones. When the zinc is 
coated with copper, the precipitation of the gold is very slow and 
imperfect. By increasing the strength of the solution to near the 
working strength, before it enters the boxes, the copper may be 
largely kept in solution. 

When the ore or tailings contain copper, the supplies of fresh 
zinc should only be added when the strong solution is passing 
through the extractor. By attending to this much of the copper 
can be kept in solution, with a correspondingly satisfactory pre- 
cipitation of the gold. 

It should, however, not be forgotten that when there is much 
copper present in the ore, it becomes imperative to allow the 
precipitation of the copper with the gold with the object of keep- 
ing the copper contents of the solutions constant. In such a case 
to keep the copper in the solutions would soon result in their 
becoming overcharged with copper and thus useless for gold 
extraction purposes. 

If the zinc turnings be placed in a solution of lead acetate, of, 
say, 10 per cent, strength, they will become covered with a porous 
coating of lead. This lead-coated zinc, by its electro-chemical 
energy, will effect the perfect precipitation of the gold, and leave 
the copper, even from the weakest solutions, unprecipitated. 
The resulting bullion obtained by this means is, however, always 
highly charged with lead. 

At the Camp Bird mills, Ouray in Colorado, the waste solutions 
are very weak, of poor value, and contain some copper. The lead 
acetate method was tried without much success and abandoned in 
favour of a mercury-couple which is obtained by immersing the 
zinc in a weak solution of mercuric cyanide until it is coated with 
mercury. The mercury coated zinc is stated to give a very fair 


precipitation, while the mercury is recovered and does not pass 
into the bullion as is the case with lead. 

Experience has shown that ores containing much copper are not 
adapted for cyanide treatment, firstly, on account of the undue 
consumption of the cyanide ; and, secondly, on account of the 
difficulty of precipitating the gold in the presence of the base 
metal ; moreover, by continued use, the stock and sump solutions 
become charged with copper, and thus rendered useless for all 
practical purposes, such as washing, or forming the basis of 
working solutions. 

Occasionally an inert gritty, greyish- white, porous precipitate 
of zinc cyanide forms on the zinc in the precipitating boxes. The 
reactions which lead to its formation have not yet been satis- 
factorily explained, but, whatever they may be, its presence is 
always accompanied by loss of cyanide and imperfect precipitation 
of the gold. This precipitate is seldom seen excepting in the treat- 
ment of pyritic ores and tailings. It can generally be prevented 
by a careful preliminary washing, and treatment with lime instead 
of caustic soda. 

On the other hand, when a too free use of lime is made to 
reduce acidity, an incrustation of lime will form on the zinc and 
thus prevent satisfactory precipitation. 

In some cases there may be, in the presence of organic com- 
pounds, an excessive and injurious evolution of hydrogen. During 
the treatment of decomposing pyritic tailings at the Great 
Mercury Cyanide Works, Kuaotunu, the evolution of hydrogen gas 
was so vigorous that it lifted the zinc out of the precipitation 
boxes, forming a thick froth. On this occasion the precipitation 
of the bullion was very imperfect and unsatisfactory, and sug- 
gested polarization. 

When a scum forms on the surface of the solution in the 
precipitation boxes, both it and its cause should be removed 
without delay. In the case of accumulated tailings it will generally 
be found to be caused by the presence of decomposing organic 
matter, and the application of an oxidizing agent often exerts a 
beneficial effect. 

In practice the zinc shavings are first placed in the weak 
solution box, and afterwards transferred to the medium, and 
thence to the strong. In the weak and medium solution 
precipitating boxes, the gold becomes plated on the zinc, and less 
zinc is destroyed than in the strong. The solution from the 
extractor boxes, containing only traces of gold, is returned to the 
solution tanks, where, if found necessary, its strength is made up 
by the addition of cyanide. 

The precipitation of the gold from very weak solutions, especi- 


ally in the presence of copper, has always been a difficulty with 
zinc, but of late years success has, to a great extent, been obtained 
by pickling the zinc, before use, in a solution of acetate of lead, 
and adding fresh cyanide at the head of the box, as was done by 
the author in 1895. 

This method of precipitation has been in use in the Lydenburg 
district for over five years, and is now largely in use at the Rand, 
to assist the precipitation of the gold contained in the extremely 
weak cyanide solutions, which necessarily accumulate in large 
volumes in the treatment of slimes. 

The following interesting particulars of the operations at the 
latter place have been given by Mr. T. L. Carter * : — 

A very important point is the preparation of the zinc. After 
it has been cut on the lathe it is taken to a trough which contains 
a solution of acetate of lead of about 10 percent, strength. The 
zinc is thoroughly washed and stirred in the solution until it 
becomes of a dark hue. If it is not thoroughly stirred only the 
outside of the mass of zinc will become coated, the inside remain- 
ing quite bright. An empirical way to ascertain whether or no 
the solution is strong enough is to look at the zinc, and see if it is 
sufficiently and thoroughly coated. After thus preparing the zinc 
it should be placed in the box, and covered with the auriferous 
solution as quickly as possible, since leaving it in the air seems 
to affect it adversely. 

The next important point is the addition of free cyanide at the 
head of the box. Twenty pounds of KCy are dissolved in an 
iron tank, holding about 75 gallons water. This 2£ per cent, 
solution is allowed to run freely into the auriferous solution 
entering the box for a period of about four hours, raising the 
strength of the solution passing through from *007 per cent, to 
•025 per cent. When the 20 lbs. of KCy are finished, another 
10 lbs. are dissolved as before, and freely run into the box, taking 
a period of six hours, and raising the solution about '007 per cent, 
higher. By no means should the addition of the cyanide as 
described be neglected, for it is found absolutely necessary at the 
commencement to let this free cyanide run into the box. Twelve 
or fourteen hours after starting a slow drip is allowed to fall into 
the solution as it enters the box, bringing up the strength of the 
solution going through the box from *007 per cent, to '008 per 
cent., and this is dropped in through the run of the box. The 
precipitation is almost perfect. On account of the lead present 
in the slimes, the smelting gave much trouble at first, but, after 
many experiments, the following flux was found to give the best 
results : — 

* Jour. Chem. and Met. Soc. S.A., No. 9. 


Borax, 60 per cent. 

in lure, . . • • r . J. «/ , , 

Sand, . . . . , . 11*5 „ 

oocia, • • . . i ,, 

The precipitation of gold by zinc, results in the formation of 
a double cyanide of zinc and. potassium, and the continual use of 
the same solutions would lead to the belief that the. working 
solutions would in time become charged with zinc salts. In 
actual practice it is found that this is not the case to any great 

Feldtmann states that, under favourable conditions, the zinc- 
potassic cyanide is of itself capable of dissolving gold from its 
ores. He considers that the small quantities of alkaline sulphides 
present in commercial cyanide, or formed by the action of cyanide 
on metallic sulphides, serve to precipitate a portion of the zinc as 
the insoluble sulphide. 

In cases where ores contain considerable proportions of metallic 
sulphides, soluble in cyanide solution, sufficient alkaline sulphide 
may be formed to precipitate a portion of the dissolved gold. 

To prevent any loss in this direction, Mr. J. S. M c Arthur 
suggested the addition of a solution of some soluble lead or other 
metallic salt which is known to form an insoluble sulphide in 
alkaline solutions, and this practice is now general when the cyanide 
contains sulphides. In this case, however, it would be advisable 
to avoid an excess of the lead, or other salt, so as to prevent 
possible complications in the extractor. The exact amount of salt 
required can be determined in the laboratory. 

The Clean-up. — The periodical clean-up takes place once or 
twice a month. The first operation is to pass a current of clean 
water through the zinc boxes, so as to remove the cyanide 
solution, which is injurious to the workmen, often causing their 
arms to become covered with painful red boils. 

The trays holding the zinc are then moved up and down in 
their compartments so as to allow the fine gold precipitates, and 
fine particles of zinc, to fall through the sieve and settle in the 
bottom of the box. The contents of the trays are then placed in 
a large trough, provided with an easily removable false bottom 
of finely perforated iron. The zinc is gently teased out and 
rubbed in this trough, which is partly tilled with clean water, 
and in this manner as much as possible of the adhering gold 
is removed. After all the gold has settled, as a slimy mass, 
the water is syphoned off, The gold slimes remaining in the 
extractor are sluiced through plugholes into the side launder, or 
into the bottom compartment and collected in a trough. The fine 


slimes or precipitates are rapidly settled by the addition of a little 
powdered alura to the solution. 

In large cyanide works the precipitates are dried in a small 
filter-press or vacuum-filter. The discoloured zinc shavings are 
now returned to the precipitation boxes, fresh zinc being put in 
the lower compartments. The gold still remaining on the zinc is 
recovered at the next clean-up. 

The gold slimes are treated in one of three ways, namely : — 

1. By smelting with suitable fluxes. 

2. By sulphuric acid method. 
3 By lead smelting method. 

The smelting method is the oldest and still has many adherents. 
It is, however, laborious, slow, and attended with loss of bullion. 
It is gradually being superseded by the sulphuric acid method, 
which in its turn seems likely to be superseded by the lead 
smelting process. 

1. Smelting Process:— Roasting the Precipitates. — The dry 
precipitates are roasted at a low heat, with free access of air. The 
object of the roasting is to oxidize the zinc in the slimes, and thus 
cause it to combine with the fluxes in the subsequent smelting, 
and thereby leave the bullion as fine as possible. 

In Australia and New Zealand cyanide works the roasting 
furnace often consists of a large flat cast-iron plate, with raised 
edges. It is built over a small grate or furnace, and a hood of 
light sheet-iron is placed over the roasting place, so as to carry off 
the zinc fumes. 

The roasting should be conducted at a moderate heat, i.e., it 
must never rise above a dull red, and the precipitates must be 
stirred continuously so as to expose fresh surfaces to the action 
of the atmospheric oxygen. During the early part of the 
roasting, dense white vapours of zinc oxide are given off, but 
as the operation advances they are observed to diminish and 
finally to cease entirely when the reaction is complete. Time, 
from one to two hours. 

Mr. Feldtmann has found that the oxidation of the zinc is 
facilitated by the addition of a little nitre, say from 3 per cent, 
to 10 per cent. He suggests that it should be applied to the 
slimes as a strong solution before the drying, so as to get 
thoroughly mixed with the whole mass. The nitre not only 
helps to oxidize the zinc, but is also said to assist the subsequent 
fluxing by uniting with the zinc oxide, and forming a ziucate of 
potash, which is not so readily reduced by the plumbago of the 
crucible as the oxide. At many works the nitre is added to the 
dried slimes in a powdered form ; of course less nitre must be 


used than is necessary to oxidize all the base metal present, as 
any free nitre remaining would rapidly destroy the plumbago 
crucible during the smelting process. Besides rendering the 
bullion fine, the nitre roasting gives a cleaner slag and greatly 
hastens the fusion. 

When stirring and removing the roasted precipitates, care 
must be exercised to avoid a loss of bullion in the form of dust. 

Smelting the Oxidized Precipitates. — The roasted precipitates 
are now placed on a large, shallow iron tray, mixed with the 
necessary fluxes, and fused in plumbago crucibles. The following 
fusing mixtures have alway given satisfactory results : — 

Clean Precipitates Much Zinc and Very 
Little Zinc. Little Sand. Sandy. 

Precipitates, . . 100 100 100 

Bicarbonate of Soda, 




6 20 50 

50 50 30 

3 15 — 

The chief essentials in a slag are fluidity, small bulk, and non- 
corrosion. The first is conferred by the " fused " borax and the 
last by the sand. 

It is desirable that the fluxes used should be free from moisture, 
so as to avoid loss due to the escape of steam through the charge. 

Smelting Acid-Treated Slimes. — Acid-treated gold slimes are 
necessarily of a basic character, hence an acid flux must be 
used so as to produce a neutral or non-corrosive slag. Manganese 
dioxide is now generally added for carrying base metals into the 
slag. Messrs. E. H. Johnson and W. A. Caldecott have shown 
that in a state of fusion it is even a more active oxidizer than 
nitre although it contains much less available oxygen.* 

The following is the basis of flux used by these authors, the 
proportions being varied, within the limits, as varying conditions 
required : — 

Slimes, .... 100 parts. 

Fused borax, . . . 20 to 35 „ 

Manganese dioxide, . . 20 „ 40 „ 

Sand, . . . . 15 „ 40 „ 

The addition of soda to an already basic material was con- 
sidered unnecessary. When sulphates are present, a little fluor- 
spar is found to assist the fusion : a few preliminary trial fusions 
should be made to ascertain the best proportions of the fluxes. 

* Jour. Chem. and Met. Soc. S.A., July 1902. 


The authors quoted above state that a matte or base-looking 
bullion indicates too little manganese dioxide, whilst too much 
manganese dioxide yields an infusible slag or one containing 
much silver. It has long been known that manganese dioxide 
in fusion carries off silver, and for this reason in smelting slimes 
containing commercial values in silver, it must be used with care, 
or not at all. 

Plumbago crucibles with removable clay-liners are generally 
used for the fusion of these charges. Crosse uses silicate of soda 
as a source of silica aud clay-lined crucibles instead of removable 
clay -liners. 

In works where large quantities of slimes have to be smelted, 
Nos. 50, 60, or 70 plumbago crucibles will be required. 

The Actual Fusion. — The crucible, previously annealed, is 
placed on a flat brick resting on the bars of the furnace. A 
priming of borax is then placed in the crucible, and over this a 
charge of precipitates ; fresh additions of precipitates are made 
'as the charge fuses and subsides. When the crucible is two-thirds 
full, the slag is skimmed off and fresh portions of precipitates 
added until it is three-fourths full of molten bullion. 

The crucible is now removed from the furnace) and the contents 
poured into ingot moulds which have been previously well heated 
and carefully oiled with the best olive oil. All excess of oil 
should be wiped out of the mould before pouring the metal. 

The melting furnace may be constructed to hold two or three 
crucibles at the same time. It should be built of the best 
materials, as the heat required to melt the slime mixture is higher 
than that for ordinary smelting. 

At the Langlaagte Cyanide Works the slimes, mixed with the 
fluxes, are charged into No. 50 plumbago crucibles, and melted in 
a reverberatory hearth furnace which holds 22 crucibles at the 
same time. The time required for melting varies from one and 
a half to three hours, according to the character of the materials 
and the temperature of the furnace. 

The slag resulting from the smelting of slimes always contains 
a small proportion of gold. It is, therefore, generally pulverized 
in a single stamper, or in a small mill, and afterwards amalga- 
mated with mercury. In some cases it is sent to the smelting 
works for treatment. 

The ingots of bullion, obtained from the first smelting, are 
re-melted with borax ; and, since gold forms but a very imperfect 
alloy with zinc, this second melting should be conducted at as low 
a temperature as possible so as to obtain an approximately 
uniform bar of bullion. 

The zinc slimes generally contain from 30 per cent, to 65 per 


cent, of bullion, the fineness of which, after melting, generally 
varies from 600 to 900. 

The clips for assaying should be taken from different parts of 
the bar so as to obtain a representative sample for valuation, but 
the dip sample is always the most reliable. 

M c Bride gives the cost of smelting gold slimes at 2d. per ounce 
of fine gold. At a clean-up giving 718 ounces of bullion, 817 
fine in gold, the costs were as follows : — 

Borax, . 

. £1 5 


. 6 


Fluor-spar, . 

. 5 


Nitre, . 

. 2 


Pots, . 

. 1 1 


Coal, . 

. 1 16 


. 1 2 


£5 19 8 = 2d. per oz. fine gold. 

2. Refining by Sulphuric Acid.— This method is commonly 

used in cyanide works in America and South Africa, but has not 
yet been adopted to any extent in Australia or New Zealand. 
The acid treatment of the precipitates is a simple enough opera- 
tion, and was occasionally used by the author for the refining of 
small parcels. The necessary apparatus consists of shallow wooden 
tubs, or vats. 

The operation is conducted as follows : — Clean water is passed 
through the zinc extractor, for some time, to remove all traces of 
cyanide. The precipitates are then removed from the boxes and 
placed in the first vat, with a sufficient quantity of dilute 
sulphuric acid. The acid should not be too strong, nor yet too 
weak ; a mixture consisting of ten parts of water and one part of 
strong acid answers well. 

The quantity of dilute acid will depend on the proportion of 
zinc present in the precipitates. With 50 per cent, of zinc, about 
six parts of the acid mixture to one of the precipitates will be 
required; and with very zincy precipitates, from ten to twelve 

The mass in the tub should be stirred occasionally, and then 
allowed to settle. Heat is generated, and large quantities of 
hydrogen gas liberated by the action of the acid on the zinc. 

When the undissolved precipitates have been allowed to settle, 
the clear liquid should be removed by decantation into a second 
tub, and thence finally, after an interval, into a third tub. By 
this means any fine particles of bullion which have escaped in the 



first decantation will be secured in the second tub ; and that 
which has escaped during the second decantation will be found 
as a fine sediment in the third tub. 

The bullion should now be washed in the tubs with clean hot 
water, to remove all base soluble sulphates and any free acid 
remaining. Then remove the bullion slimes and dry on the 
vacuum-filter. When dry, subject to an oxidizing roasting on a 
shallow iron pan, for an hour or so, to oxidize any base sulphates 

Next, mix with 5 or 10 per cent, of borax glass, according to 
the amount of zinc oxide still present, and fuse in a plumbago 
crucible in which a priming of borax has already been placed. 
As the charge fuses and subsides, fresh portions of bullion should 
be added, until the crucible is three parts full of melted metal. 
To permit of this being done the slag can be skimmed off from 
time to time. 

If manganese dioxide is used in the flux it will be necessary to 
protect the crucible from corrosion by using a clay-liner as 
described in the preceding pages under the heading devoted to 
the " Smelting of acid-treated Slimes." 

The bullion is generally from 850 to 900 fine, but with a little 
extra trouble can be worked up to 950. 

With suitable appliances, this process possesses many advan- 
tages over the smelting process. It occupies less time, produces 
finer bullion, and, properly conducted, costs less. 

When large quantities of precipitates have to be dealt with, the 
method of settling the slimes and decanting is too slow and expen- 
sive. In this case, the separation of the slimes from the acid 
solution, as well as the subsequent washings, must be effected in a 
vacuum-filter, or a filter-press, as practised in America and South 

The filter used for the purpose is a wooden box, two or three 
feet square. It is provided with a filter webbing, or cloth, of fine 
canvas or twill duck, resting on a grating of wood, and fixed with 
slips of wood, so as to be easily detached for washing. 

The false bottom, below the webbing, must be 15 in. or 20 in. 
deep, and provided with a solution gauge, the upper limb of which 
should be 2 in. below the air-exhaust pipe connected with the 
vacuum boiler. 

Care must be taken to draw off the acid solution and washings 
by a plug-hole before they rise to the level of the air-exhaust 
pipe, which is placed immediately before the filter -frame. 

When the acid solution is diluted to half its strength before 
filtering, the webbing lasts for several operations. A Johnston 
filter-press does the washing well and expeditiously. 


Acid Treatment in South Africa. — The following description is 
an abstract of a paper read by E. H. Johnson before the Chemical 
and Metallurgical Society of South Africa. It describes the 
process employed at the Princess Works, where the slimes are 
submitted to acid treatment before smelting, and will be especially 
interesting for its figures of costs and metallurgical results : — 

At the Princess Works the slimes from the zinc boxes are 
separated from the solution drawn off with them by the aid of 
a vacuum-filter. A water wash is passed through until the slimes 
are free from cyanide. The gross weight of the slimes, including 
moisture, is then taken, by weighing the buckets of moist slime 
during transference to a large sheet-iron tray placed alongside the 
acid tank, to determine the amount of sulphuric acid necessary to 
destroy the zinc. 

Having found the approximate weight of slimes to be treated, 
sufficient water is run into the acid vat to form, on the addition 
of the acid, a 10 per cent, solution. One pound of acid for every 
pound of moist slime gives good results. This would be equivalent 
to about 1 J lb. of acid to the pound of slimes, dry. The acid is 
then added, and the vat closed down tightly. 

The stirring apparatus is kept continually moving during the 
time of feeding in the slimes, which are fed in gradually in the 
same condition in which they were taken from the filter-vat. It is 
beneficial to keep up a continual stirring for at least half an hour 
after the action has apparently ceased. 

After all the slimes are in the acid, a jet of water is turned into 
the hopper to wash down any adherent slimes, and everything 
that has been used in the cleaning-up of the boxes, etc., is well 
washed in the same jet during removal. The vat is then filled 
with water and allowed to settle. 

Working with dilute acid, and not heating, a perfect settlement 
takes place within an hour. When heating with a steam jet, 
settlement was much more difficult. 

The washing is done by syphoning off the clear liquor, and 
filling the vat repeatedly with water, until the solution is neutral 
to litmus paper — usually four or five washings. It is well stirred 
at each refilling by means of a long wooden paddle, a rotary motion 
being given to the water. This causes the slimes to collect in the 
middle of the vat, and reduces the risk of loss during syphoning — 
the syphon being let down at the side. A sample of the washings 
taken continually during syphoning off showed, on careful assay 
of a large sample, 13 gr. of gold per ton of solution. 

The drying of the resultant gold slime is conducted on an open 
drying hearth in large cast-iron enamelled dishes. The cakes are 
subsequently broken up and transferred to small sheet-iron trays 


in thin layers, and subjected to an increased heat. When cool 
the slimes are ground, fluxed, and transferred to the crucible. It 
fuses quietly and with but little fume, and normally yields 50 to 
60 per cent, of the weight of slime as bullion. 

The average fineness of last year's bullion was, according to the 
Work's assays, 821*9 and 819*6, according to London returns. As 
a deduction of 2 milliemes is made on 800 bullion, this leaves an 
actual difference with London of 0*3 milliemes. 

The slag, after panning out the prills (of which there is very 
little), assays 23 oz. per ton, and one ton of slag has been accumu- 
lated in two years for an output of 11,627 oz. of fine gold, 
which is equivalent to a little under 0*2 per cent, of the total 
gold produced. 

The cost of reduction, including acid, is 6'7d. per fine ounce, 
made up as follows, taking an actual "clean-up "as a basis: — 
Dry weight of zinc-gold slimes, 504 lbs. ; dry weight after acid 
treatment, 100 lbs. ; 672 lbs. of acid at 4£d., £12, 12s. lOd. ; 66 lbs. 
of borax at 37s. 6d. per cwt., £1, 2s. Id. ; 9 lbs. carbonate of soda 
at 2£d. per lb., Is. lOd. ; 9 lbs. fluor-spar at 4d. per lb., 3s. ; 5 bags 
coke at 8s. 6d. per bag, £2, 2s. 6d. ; 1 No. 60 crucible, £1, 7s. 6d. ; 
total, £17, 9s. 9d. ; yield, 620 oz. fine gold, or 6*7d. per fine oz. 

From information gathered from reliable sources Mr. P. S. 
Tavener states that the cost of sulphuric acid refinement of slimes 
in South Africa is as follows * : — 

a. Sulphuric acid treatment from 4d. to 8d. per oz. of fine 

b. Smelting cost exclusive of labour and furnace wear and 
tear 4d. to 6d. 

Approximately this gives a total cost of Is. per oz. of fine gold 

Acid Treatment in America. — I am indebted to Mr. Godfrey 
Doveton, one of my former assistants, for the following par- 
ticulars of the acid treatment at the Camp Bird Mills, Ouray, 
in Colorado : — 

" The gold slimes are refined by sulphuric acid in a wooden vat 
12 ft. in diameter and 2 ft. deep, coated with paraffin paint. 
They are not subjected to heat, sufficient being generated on the 
addition of the acid to the moist slimes. About twelve hours are 
occupied in dissolving the zinc, and some forty-eight hours in wash- 
ing out the soluble sulphates. The washing is much expedited by 
the use of hot water, at least for the first six washings. About 
fifteen wash waters are usually required. After the first two 
washes, having a high specific gravity, are removed, the time 

* Tavener, Jowr. Chem. and Met. Soc. S.A., Oct. 1902. 



allowed for settling is one hour. Usually it takes from eight to 
twelve houi*s to remove the first two washes, as the slimes do not 
settle so readily in the heavy acid liquid. 

" All the washings are stored in a large settling-tank, which is 
cleaned up at long intervals. The settling-tank, after a year's 
run, yielded bullion worth £20. 

" The first two or three wash waters on assay are found to yield 
from 10 to 20 cents per ton, but subsequent washes contain from 
5 cents to a trace, — only apparent on assay of large evaporations. 

" The gold slimes are partially dried by filter-press, and are then 
transferred to a hearth calcining furnace and calcined. The 
hearth of furnace has an area of 36 square feet. The furnace-doors 
are securely padlocked, and the furnace serves the double purpose 
of a safe and calciner. 

"From the furnace the slimes are removed as required, and 
charged into graphite crucibles, and smelted, having been pre- 
viously mixed with 50 per cent, of a flux composed of the following 
parts: — 

Soda, .... 2 parts. 

Sand, 1*5 „ 

Borax powder, . . 4 „ 

Sulphur, .... as required. 

Thus 200 lbs. of slimes require 100 lbs. of flux. The charge is 
smelted with frequent skimming in from ten to sixteen hours for 
each crucible. 

" The slimes contain a considerable quantity of copper, but com- 
paratively little zinc, consequently little or no nitre is used ; but 
we use instead a certain quantity of flowers of sulphur, which 
readily converts the copper to a matte or regulus, leaving the 
bullion comparatively fine. The regulus, which by the way is 
usually from 15 to 25 per cent, of the weight of the bullion, was 
found to contain, on assay — 

Gold, . . . 23*4 oz. per ton. 

Silver, . . . 408*0 

Copper, . . . 60% 

Zinc, ... 3% 

"The slag resulting from the bullion melting is ruby red in 
colour, and contains — 

Gold, . . 15 oz. ) Assayed without removing 

Silver, . 265 „ J shots or prills of bullion. 


Copper, . . 6 to 8% 

Zinc, . . 12 „ 16% 


"The bullion averages about 800 to 850 fine, the impurity 
almost entirely copper with a little zinc. 

" A shipment of 3000 lbs. of slag and regulus mixed was lately 
sent to the smelter and netted us £120." 

3. Tavener Lead-Smelting Method. — This process was 

first made use of by Mr. P. S. Tavener in August 1899. Since 
August 1901 it has been in continuous operation at the Bonanza 
Mines, Limited, Johannesburg, and already the process has been 
adopted by many leading mines at Johannesburg. The adoption 
of the lead-smelting of gold slimes marks a notable advance in 
cyanide practice*. The old smelting process is laborious, and 
always likely to entail serious losses of gold where large quantities 
of slimes have to be handled, while the sulphuric acid method is 
cumbersome, slow, and costly. 

The obvious advantages of lead-smelting compared with the 
sulphuric acid process may be briefly summarized as follows : — 

a. Saving of cost per oz. of fine gold produced. 

b. No by-products. 

c. Less liability to loss in handling slimes. 

d. More gold actually produced from a given weight of slimes. 

The essence of every metallurgical process is its cost, and judged 
by this principle, lead-smelting possesses a marked advantage. In 
South Africa, where the sulphuric acid method has, perhaps, its 
greatest application, the average cost, according to Tavener, is not 
less than one shilling per oz. of fine gold produced. The lead- 
smelting process costs threepence per oz. of fine gold, including all 
charges, which means a saving of ninepence per oz. compared with 
sulphuric acid method. 

Thus the lead method at threepence per oz., in a mine producing 
2500 oz. fine gold per month from cyanide works, would effect a 
saving of costs amounting to £93, 15s. per month, equal to about 
£1000 a year. 

The Tavener lead-smelting process is cheap, rapid, and efficient 
compared with other processes, and with some modifications in 
practice is likely to be universally adopted. 

The following working details of lead-smelting as practised at 
the Bonanza mine are extracted from a paper read by Mr. Tavener 
before the Chemical and Metallurgical Society of South Africa last 

Mr. Tavener says that he can best describe lead-smelting by 
comparing it to a scorification assay conducted on a large scale, 
for the reason that the zinc slimes are melted and the gold 

* Jour. Chem. and Met. Soc. S.A., Oct. 1902. 


recovered in lead bullion. The lead bullion is then cupelled, or 
to use a better term, refined. 

"The * clean up,'" he continues, "is conducted in the ordinary 
way, with the exception that all the precipitate is at once pumped 
from the i clean-up ' tub into the filter-press. The fine zinc which 
remains at the bottom of the ' clean-up ' tub is heaped up on one 
side and allowed to drain for about half an hour, and is then ready 
for the smelting room. The filter-press is cleaned out and the 
cakes taken in their moist condition to the furnace, and there both 
slimes and fine zinc are put in trays into a drying oven, and 
sufficient time is allowed to warm through, fifteen minutes for 
each tray being sufficient. Care is taken to keep the fine zinc 
separate from the filter-press slime, and on no account should they 
be allowed to get mixed. 

"In charging the furnace, the slime is first dealt with. After 
warming in the drying oven, it is at once rubbed through a sieve, 
four holes to the linear inch, and then roughly weighed for fluxing, 
the necessary fluxes having been previously mixed. 

" The slime is mixed with the fluxes and passed through a sieve 
to ensure thorough mixing. It is then shovelled direct into the 
furnace. When all the filter-press slime is fluxed and charged, the 
fine zinc is dealt with in the same way, and put into the furnace 
on the top of the slime in order to prevent loss by dusting, and 
also to have the greater portion of litharge present on the top of 
the charge. The fluxes used are residue assay slag and commercial 
litharge. The former costs nothing, for unless used in this manner 
it would be thrown away." 

Discussing the question of fluxes, he says : — " 1 have found that 
the following, with little variation, will give satisfactory fusion 
and clean slag : — Slag, 25 to 30 per cent., made up of 10 per cent, 
assay slag ; the balance, equal quantities of old slag and scalings 
from the pots of previous smelt. In the event of the lead-smelting 
method being adopted, I should like here to point out the advan- 
tage of storing the slag now being obtained from crucible smelting, 
since in the lead process its gold contents are converted into bullion 
free of cost. If it were not treated, clean slag would have to be 
used. During the last year I have been able to deal with several 
tons of this material assaying over fifty ounces fine per ton, left 
behind on the mine after the Boer occupation. The same applies 
to clay liners and anything else carrying gold, now termed by- 
products. With a lead-smelting furnace the word by-product is 
forgotten, since none is obtained." 

The quantity of litharge to be used will depend on (a) value of 
slimes, (b) weight of charge in furnace, and (c) the percentage of 
gold required in the resultant lead bullion. 


Mr. Tavener gives the following proportions of fluxes for gold 
slimes : — 

Gold slimes, . . . 100 parts by weight. 

Litharge, ... 60 „ 

Assay slag, . . 10 to 15 „ 

Slag previously used, . 10 to 15 „ 

Sand (Si0 2 ), . . . 5 to 10 „ 

Sawdust, . . . 1% of weight of litharge. 

For fine zinc he uses the following proportions : — 

Fine zine, . . 100 parts by weight. 

Litharge, . . . 150 „ 

Slag, .... 20 „ 

The products of different mines will necessarily vary, and the 
proportions to give a clean, well-oxidized slag can easily be deter- 
mined by experiment on a small scale with assay crucible tests. 
Mr. Tavener, however, mentions that considerably less assay, or 
other slag, will effect a good fusion in a reverberating furnace than 
in a crucible, and he states that it would be safe to use 30 per 
cent, less slag when smelting in the reverberatory than was found 
necessary in the crucible trial. 

The quantity of litharge should be so proportioned that the lead 
bullion should not carry more than 8 per cent, of gold, or 10 per 
cent, at the maximum. It was found preferable to make a larger 
quantity of lead than to have it too rich in gold contents. 

No reducer is used with the fine zinc, which is relied on to 
reduce sufficient lead, leaving an excess of litharge in the slag to 
ensure success. With the gold slimes charge, 1 per cent, of saw- 
dust is added on the weight of litharge, but if a larger proportion 
of litharge has been used, then from 1£ to 2 per cent, of sawdust 
is necessary. 

The furnace work is described as follows : — 

"When the entire charge of zinc and slime is in the furnace, it 
is banked up from the sides to the centre so as to avoid the 
possibility of particles remaining on the sides above the slag level 
as the charge reduces and settles down. A covering of litharge is 
spread over the surface, and on this again a light covering of easily 
fusible slag is spread. The furnace is charged the day previous 
to smelting, and one of the night-shift men lights a slow fire 
about 3 a.m., which serves to dry the charge. At 5 a.m. 
the damper is opened and the fire urged, and in half an hour 
the furnace is at a smelting heat. By 9 or 10 a.m. the charge is 
reduced, then sweepings from cyanide works, smelting room, or 


any slag requiring re-smelting, is added and is quickly absorbed in 
the molten bath. When all this has been fed in and melted, and 
the slag become fluid, it is well stirred with a rabble, and sawdust 
is thrown in to reduce the excess of litharge in slag. This opera- 
tion is repeated until the slag, which remains on the rabble when 
withdrawn from the furnace, is judged by its appearance to be 
olean. The slag is now run off into pots through the slag-door, 
the level of which is 4 in. above the centre of the lead bath — a 
bath of 12,000 oz. of lead bullion almost occupies this space. 

"Before filling the furnace, the slag-door is built up about 
12 in. by placing flat cast-iron plates, £ in. thick, bedded in fire- 
clay, one on top of another, and in front of these plates a bank of 
fire-clay is also made. In order to run the slag off, all that is 
necessary is to break away this bank, plate by plate, and so allow 
the slag to flow over into the pot. When the pot is full it is 
wheeled away and another is put in its place. The filled pot is 
run outside, and after standing a minute or two, is tapped, and 
the molten slag allowed to run out on the ground to cool ; that 
which remains on the sides and the bottom of pot is brought back 
for further use. When no more slag will flow from the furnaces 
owing to the bath being down to the level of the slag-door, it is 
waved off by rabbling. At first sight it would appear difficult to 
draw this remaining slag off without dragging out some lead, but 
a very little practice enables it to be done so closely that there is 
little but a thin skimming of slag remaining. In the event of a 
little lead being pulled out into the pot, it is recovered from slag 
pots. It is for this reason that the pots are tapped about 2 in. 
from the bottom. By opening the fire door this last skim on the 
lead bath quickly thickens. A shovelful of lime is thrown in to 
assist. This skim is easily pulled off, and of course is held over 
until next smelt. By this means a clean surface of lead is exposed, 
and any zinc present would be quickly got rid of, for at this stage 
the lead is at a bright red heat, and the free access of air due to 
the open fire door quickly oxidizes it. So far, lead recovered by 
this method has always been clean and soft, a proof that no zinc 
could be present, since one per cent, of zinc gives lead a distinct 
silvery colour, and makes it so hard that it cannot be rolled. The 
lead bullion is tapped by driving a \ in. steel bar, tapered to a 
point, into the tap-hole, which is closed with a fire-clay plug. The 
lead is run into an iron trough, which conveys it to the moulds 
placed together on the floor." 

Before tapping the furnace the lead-bath is well stirred, and a 
sample is taken out with a ladle and granulated. 

The cupelling or refining of the lead bullion is next described, 
with particular details and some useful hints on the making of the 


bone-ash test and regulation of blast. The process differs only in 
minor details from the usual operation of lead-refining, and need 
not be specialized in this work. 

The costs at the Bonanza, Limited, for stores for cupelling and 
smelting for four months (June-Sept. 1902) were as follows : — 

Coal, . 

£82 13 


Coke, . 

Fire-bricks and slabs, 
Paper bags, 
Lead foil, 

9 8 
17 15 
1 17 




Bar iron, 



Crucibles and liners, 

1 15 

Caustic potash, 




1 5 


Cost per fine oz 

Lead loss estimated at 
12% on 15,269 lbs. = 
1832 lbs., say 1 ton at 
£20, . 

£131 10 7 = 2-463d. 

20 = 373d. 

£151 10 7 = 2-837d. 

Gold output in period, 12,810 oz., fine. 

It is claimed that lead-smelting recovers a larger amount of gold 
than the sulphuric acid method, and this claim seems to be justified 
by several large experiments on equal weights of identical gold 
slimes by both processes. In six trials the lead process gave 10*5 
per cent, higher recovery than the acid treatment. These results 
are probably much higher than could be obtained in continuous 
working ; but it is evident that if only one or two per cent, more 
gold can be recovered, the lead-process has fully established the 
claims of Mr. Tavener. 

In the discussion which followed the reading of Mr. Tavener's 
paper, Mr. D. J. Williams suggested that the probable loss of gold 
by volatilization in the cupellation, caused by the presence of such 
a volatile metal as zinc, might be avoided by first dissolving the 
zinc-gold slimes in sulphuric acid, washing, drying, and then sub- 
jecting to lead-smelting. Mr. Tavener agreed with this proposed 
modification, and stated that he would rather receive at the furnace 
acid-treated precipitate, which would render the work easier and 
less troublesome. He further stated that with zinc present, a 


certain amount of experience and skill was required, while with 
acid-treated material little or no experience was needed.* 

That lead-smelting recovers a higher percentage of gold than 
the acid-treatment seems to be established on pretty conclusive 
evidence, but it is not quite so easy to point to the source of gold 
thus recovered; that is, to show where the loss takes place in 

Zinc Dust Precipitation. — The method of application is to 
agitate a certain quantity of the dust with the gold-containing 
cyanide solution in vats, allow the precipitate to settle, and 
decant the clear solution. 

During the past year or two this method of precipitation has 
been adopted at a number of American mills, among which may 
be mentioned the Homestake Mining Company, South Dakota ; 
the Montana Mining Company, Montana ; the De La Mar Mining 
Company, of Nevada ; and the Consolidated Mercur Company, of 

At the Homestake Company's cyanide works u Precipitation is 
carried on by means of zinc dust in five tanks 25 ft. in diameter 
and 20 ft. deep, built of California red-wood. The bottoms are 
inclined to one side, where a small sump is provided in order to 
drain the tanks perfectly. No gold storage solution tanks are used, 
the solutions draining from the leaching vats directly into the 
precipitating vats. The precipitation vats are pumped out by 
duplex Prescott pumps, the solution going to filter presses made 
in the Homestake shops. The pulp treated by the mill is of a 
highly siliceous nature, often however carrying some hornblende, 
the sulphides being mainly pyrite and pyrrhotite The extraction 
made is in the neighbourhood of 90 per cent., the cost being 
between 45 and 50 cents per ton of material treated, probably 
nearer the latter figure."! 

In America the price of zinc dust varies from 25s. to 29s. per 
cwt., and the amount of dust used for precipitation varies from 
6 to 9 oz. per ton of solution. 

Charcoal Precipitation.— At many cyanide plants in Vic- 
toria charcoal is being used to precipitate the gold from cyanide 
solutions. The solution is passed through a series of barrels 
packed with finely broken charcoal, on which the gold is deposited. 
The charcoal is afterwards burnt to an ash, and the ash fluxed. 
The process is too slow and cumbersome to recommend itself for 
use in large plants where hundreds of tons of solution have to be 
handled in the twenty-four hours. 

* Jour. Chem. and Met. Soc. S.A., Jan. 1903. 

t C. H. Fulton, The Engineering and Mining Journal, June 4, 1902. 





At the Witwatersrand Goldfields the cyanide process has been 
conducted on a more extensive scale than elsewhere. The ore 
there is principally a pyritic, silicious quartz-conglomerate, con- 
sisting of rounded or sub-angular pebbles of bluish-grey quartz 
embedded in a quartzose matrix. The pyrites occur in varying 
proportions in different mines, but the average is probably not 
less than 2 per cent. The gold does not exist in the quartz 
pebbles, but occurs disseminated throughout the matrix or in the 
iron pyrites. 

This ore is locally known as "banket," or almond rock. 
It is comparatively hard and somewhat splintery, and often 
contains a small proportion of corundum and clay, which renders 
it tough and hard, and forms slimy products during the 

Throughout these goldfields the universal practice at present 
is wet-crushing with Californian stamps, copper-plate amalgama- 
tion, concentration or classification of tailings, cyanide treat- 
ment of tailings, and chlorination or cyanide treatment of 

At most of the batteries a 30-mesh screen is used, but in a few 
cases a finer or coarser mesh is adopted. 

The main features of the cyanide treatment at the Witwaters- 
rand Goldfields are much the same as those practised in America 
and Australia. The general working details are given below in 
tabulated form. They are the same at all the cyanide plants, with 
minor differences according to the individual fancy of the chemist 
or metallurgist in charge of the operations, and the requirements 

of the ore or tailings. 




Filling vats, 12 

Preliminary alkali or water wash, if necessary, . 12 
Strong solution, 0'2% to 0'4% solution, £ to J of 

ore — 

In contact with tailings, . . . 12 

Percolating, 12 

Dry or air- leaching, 4 

Weak solution, 0-15% to 0*2% solution, about \ 

of ore, ........ 12 

Dry or air-leaching, 4 

Two weak cyanide washes, 0*05% to 0*1% washes, 

each about \ of ore, 12 

Two clean water washes, each about £ of ore, . 12 

Discharging vat, 8 

Total, ... 100 

The total quantity of solution used, including the water washes, 
is about equal to the weight of the ore. The quantity of strong 
solution used varies according as a preliminary washing with a 
dilute solution has been employed or not. In the former case it is 
about 25 per cent, of the weight of the ore, and in the latter case 
about 40 per cent. 

The percolation vats are charged with tailings to within a few 
inches of the top, and their surface is levelled. The strong cyanide 
solution is then allowed to penetrate the tailings until they are 
covered. The contents of the vat settle some inches, the amount 
of shrinkage depending on the depth of the vat and the percentage 
of moisture in the tailings. 

The value of the tailings varies from 12s. to 20s. per ton, and 
the actual extraction amounts to 70 or 75 per cent., at a cost 
varying from 4s. to 10s. per ton, according to the size of the 

The Jumpers Deep cyanide plant and slime works are among 
the newest on the Hand, and represent the most advanced and 
up-to-date practice. The cyanide works of the Waihi Company at 
their new mill at Waikino are modelled on the latest South 
African practice and experience ; and as they are fully described 
in the next pages, it will be unnecessary to give further details 
of South African works and practice, all of which are lucidly 
described by Mr. John Yates in his recent work on metallurgical 
engineering on the Rand.* 

* John Yates, " Present Day Metallurgical Engineering on the Rand/' 1898, 
The Mining Journal, London. 


New Kleinfontein Mine. — In a paper read before the Institution 
of Mining and Metallurgy, London,* Mr. F. Card well Pengilly 
gives some interesting details of the successful treatment of tailings 
by the "direct filling process." The plant, he says, consists of 
2 slime separators, 19 treatment vats, 4 extraction boxes, and 9 
solution tanks. The slime separators are of the ordinary spitz- 
kasten or pointed box style, 6 ft. square at the top and 6 ft. 
deep. Of the 19 treatment vats, 13 are of 200 tons capacity, and 
6 of 130 tons capacity each. Around the top of each vat is a 
launder, which carries away the overflow into the slimes race. 
The 9 solution tanks have a total capacity of 636 tons of solution, 
and each is connected with the three centrifugal pumps worked 
from a motor that pumps the weak, medium, or strong solutions 
on to the treatment vats. The pulp or tailings, after leaving the 
battery plates, is lifted by means of a tailings wheel into a launder, 
by which it is conveyed to the slime separators. The slimes from 
the first separator pass into the second, whereby a quantity of fine 
sand is saved that otherwise would flow away. About 15 per cent, 
of slimes are eliminated in the separators, and the remaining pulp 
is run through launders into the treatment vats, which are pre- 
viously filled with water. In the course of filling the vat, another 
10 per cent, of slimes overflows into the slimes race, so that the 
resulting tailings to be treated in the vat contain but a small 
proportion of slimes. 

The vat is filled with tailings (sand) to within 1 foot of the top, 
when the stream is diverted into another vat. 

Method of Treatment. — Each vat, after being drained of water, 
is treated with an alkaline solution to neutralize the free acid in 
the tailings formed by the decomposition of pyritic ores. As soon 
as the solution draining away is slightly alkaline, the treatment 
by cyanide solution is commenced. Various strengths of solution 
are pumped into the vat, each vat receiving during its course of 
treatment from 200 to 275 tons of cyanide solution. The length 
of treatment is six days, and the amount of cyanide used is 6 lbs. 
per ton of tailings. 

The following table shows the exact method of procedure adopted 
in the works in the treatment of a vat of tailings : — 

* Trans. Inst. Min. and Met. London, vol. vi. p. 113, 1898. 



Vat I. Charge I. Capacity 207 Tons. 

Date filling. 







Wed. 6th 

1 p.m. 

■ • • 


) 9 

2 a.m. 

• • • 

Filled leaching dry. 

• • 

4 ,, 


Alkaline wash, 20 lb. ; NaHO 

1 * 

W 9 

used, 20 lb. ; \ bag-lime. 



9 „ 


Preliminary solution, 0*16; 
KCy, 20 tons. 



2 p.m. 

• • • 

0*01 changed to weak box. 


2 „ 

• • • 

1st strong, 30 tons 0*3 KCy; 3 
p.m., 20 tons; 7 p.m., 6 



9 ,, 


Standing under solution for 3 


12 p.m. 

• • • 

Started leaching until 12 noon. 

Thurs. 7th 

1 a.m. 



Changed to medium box, 3 
a.m., 1'3; 6 a.m., 1*6; 9 
a.m., 0*18. 


12 noon 

1 0*2 

Changed to strong precipi- 
tating box. 


12 „ 

• • • 

2" strong, 30 tons ; 0*3 KCy, 
3 p.m., 10 tons ; 6 p.m., 15 
tons ; 8 p. m. , 1 ton. 



8 p.m. 


Standing under solution for 3 
hours. Started leaching 
until 8 a.m. 

Frid. 8th 

8 a.m. 


Medium, 20 tons ; 12 noon, 5 
tons; 4 p.m., 8 tons; 8 
p.m., 8 tons. 



9 p.m. 


Changed to medium ; 12 p.m. , 
8 tons ; 5 a.m., 10 tons ; 12 
noon, 7 tons. 


Sat. 9th 

3 „ 


4 p.m., 10 tons; 8 p.m., 7 
tons; 4 a.m., 12 tons; 8 
a.m., 5 tons. 


San. 10th 

6 „ 


Weak, 5 tons; 12 p.m., 5 
tons; 6 a.m., 5 tons; 12 
noon, 5 tons. 


Mon. 11th 

6 a.m. 


Changed over to weak. 


6 p.m. 


5 tons ; 12 p.m., 5 tons ; 7 
p.m., 5 tons ; 12 p.m., 4 


Tues. 12th 

6 a.m. 

• • ■ 

Leaching dry. 

Wed. 13th 

6 „ 

• • • 



Amount of KCy used, 5*9 lbs. per ton. 


Precipitation of Gold in the Extractor Boxes. — The precipita- 
tion of gold from the solution coming from the treatment vats is 
effected, he says, by the means of zinc shavings. A certain portion 
of the ore coming from the mine is of a rebellious nature, containing 
sufficient foreign metals to retard, if not prevent, the precipitation 
of gold by zinc. To counteract the effects of these foreign metals, 
it has been proved efficient in practice on these fields to set up a 
lead couple, and at these works this is effected by dipping the zinc 
shavings in a weak solution of acetate of lead, preparatory to being 
placed in the extractor boxes. 

In all cases it is found that zinc containing from 1 to 2 per cent, 
of lead gives the most satisfactory precipitation. By the use of 
acetate of lead a perfect precipitation is brought about; only 
traces of gold remaining in the solution after passing through the 
extractor boxes. 

The actual working costs amount to Is. lid. per ton, including 
general charges and maintenance. 

It has been shown in America and Western Australia that 
with certain classes of ore and certain local conditions, it may be 
more advantageous to dry-crush than wet-crush preparatory to 
treatment. In South Africa wet-crushing is universal. Dry- 
crushing in that country has never been viewed with much 
favour, nevertheless the experiments made by Mr. Franklin 
White at the Luipaards Wei Estate mine are interesting and 

The tanks used were made of steel, 25 ft. in diameter and 
8 ft. deep, with the usual filter bottoms and bottom discharge 
doors. Three tanks were placed close to the mill and mine in a 
row some distance away, and a little below. The ore was treated 
a few days in the upper tanks and then transferred to the lower, a 
double treatment being thus obtained. Mr. White does not con- 
sider double treatment to be of advantage in dealing with dry- 
crushed material, as the ore is properly mixed in the first instance, 
and there is abundance of air entangled in the dry sand. He is 
rather of the opinion that when the damp ore is transferred to the 
second row of tanks, there is a tendency to pack closer and to 
retard filtration. 

Lime was added to the ore at the rock-breaker floor, in propor- 
tions varying from 1 lb. to 2£ lbs. per ton. By this means it was 
thoroughly mixed in the different machines, and any lumps were 
broken up. 

The actual cyanide treatment differed a little from the ordinary 
procedure, and followed lines first suggested to the author by Mr. 
W. K. Feldtmann. 

* Trans. Inst. Min. and Met. London, vol. vii. p. 124, 1899. 


As soon as the tank in the first row was filled, a strong 
solution, 0'25 per cent. KCy, followed by two others, the last 
being 0*15 per cent., was pumped on to it. The time given to 
this treatment would be sixty-six to seventy hours. The solutions 
were not allowed to stand, but were drained off when the tank 
was once filled. The object of this was to allow fresh air to 
obtain access to the mass of damp sand. Each solution would be 
about 27 tons to the tank of ore (165 to 170 tons). 

The ore was then transferred to the second row of tanks, 
having lost about 67 per cent, of its original assay value in this 
short time. It is probable that the abundance of air entangled in 
the dry sand, as compared with what would be held in a tank of 
sand settled with water, materially assisted the solution of the gold. 
Also the finer grains of the free gold would be quickly taken up. 

A solution of 20 to 25 tons, not exceeding 0*20 per cent. KCy, 
was then pumped on to the transferred sands and drained off, the 
mass being allowed to remain damp for about ninety-six hours, 
when weaker solutions (0*15 per cent, to 0*10 per cent.) were 
used in continuous washes, making up a total of 75 tons per tank 
(second treatment). A water wash of 20 or 30 tons completed 
this part of the process, which would last some 275 hours. 

A careful series of moisture tests and measurements of solution 
sumps during the treatment of four tanks (680 tons) showed that 
the total loss of liquid in the treatment was 96 tons, or 24 tons 
per tank. The moisture in the discharged residues averaged 12*3 
per cent., or, say, 20 tons per tank; the remaining 4 tons would 
be represented by evaporation from surfaces of sumps and tanks 
and by leakages. As the ore contained about 3 tons of water per 
tank, in the form of moisture when delivered from the mill 
storage bin, the actual consumption of fresh liquid equals 21 tons 
per tank, or £ ton (25 gallons) per ton of ore. 

The solutions running from the first row of tanks carried from 
13 dwt. to 32 dwt. per ton ; those from the second row 3 dwt. to 
4 dwt. ; and the final wash 0*8 dwt. to 2£ dwt. 

Summary of Cyanide Costs. 

Trial Crushing (Coarse). , 

Tanks and extractor house work, . per ton 2 3*30 

Clean up and smelting, etc., 
Discharging residues, 






3 5-86 

Actual extraction 68*91 per cwt. 



The principal gold-bearing formation of the Hauraki Goldfields 
is of volcanic origin, consisting of a great accumulation of andesitic 
lavas, tuffs, breccias, and agglomerates of lower tertiary age. 
These rocks everywhere bear evidence of having been subjected 
to the prolonged leaching action of thermal waters, doubtless 
accompanied by steam and acid vapours. They are found in all 
stages of decomposition or alteration; and in many mines the 
hard blue andesite can be seen to pass by a series of almost insen- 
sible gradations into a soft, or fairly hard, greyish-yellow or blue 
altered rock, to which the distinctive name propylite has been 

It is in this altered andesite that the veins yielding payable ore 
occur. The veins vary from a few inches to 40 ft. in width, 
but in linear extension they can seldom be traced for any consider- 
able distance. The ore values are also irregular, and in no case, as 
yet, is the same vein or lode worked with payable results in two 
adjoining mines. 

Cyaniding Ores, — In* the southern portion of the Hauraki penin- 
sula, the pay-ores consist of whitish-grey chalcedonic or crypto- 
crystalline quartz, often possessing a wavy, banded structure of 
alternating layers of grey and blue flinty quartz. They are com- 
paratively free from base sulphides. 

The gold is about *645 fine, and usually associated with silver 
sub-sulphide (Ag 2 S) in varying proportions. It is generally ex- 
tremely finely divided, being seldom visible to the eye, and in the 
great bulk of the Waihi ore it is impossible to raise even a colour 
by panning. 

. Prior to the introduction of the cyanide process, these ores were 
treated by dry-crushing and hot pan-amalgamation with chemicals, 
by which a recovery of 65 per cent, was effected. 

When cyanide treatment was adopted, dry-crushing was naturally 
continued at the different mills, the dry pulverized material being 
charged into shallow vats and treated directly by cyanide. From 
65 per cent, by pan-amalgamation, the recovery rose at a bound to 
85 per cent., and in some cases to 90 per cent., and the results 
were considered so satisfactory that no further improvement was 
considered possible. 

In a few years, however, it became apparent that dry-crushing 
possessed many disadvantages compared with wet-crushing, the 
principal being the cost of drying the ore, the low duty of the 

* Excerpt from paper by author read at California Meeting, September 1899, 
American Institute of Mining Engineers. 


stamps, and the large number of vats required for leaching. In 
1897 mine owners began to turn their attention to wet-crushing, 
and one by one, since the beginning of 1898, the different mills 
have been adopting wet-crushing, until at the present time dry- 
crushing is the exception and not the rule, as it was two years 

Wet-crushing Practice. — A. For ores containing a large pro- 
portion of free, easily amalgam able gold, with a small proportion 
of fine or "float gold " and silver sulphide, the mill practice is : — 

(a.) Crushing with water. 
(b.) Plate amalgamation, 
(c.) Spitzlute separation of sands and slimes. 
(d.) Cyanide treatment of sands and slimes by ordinary per- 

A typical example of an ore of this class is that of the Kauri 
Gold Estates at Opitonui, where a new 40-stamp mill has just 
started. The sands and heavy slimes are subjected to the 
" double " cyanide treatment, but it is doubtful if the additional 
saving will pay for the extra labour involved. So far no provision 
has been made for the treatment of the fine slimes. If they are 
worth it, they will probably be treated by agitation and decanting. 

B. For a clean ore, almost identical with that described above, 
but containing a small proportion of free amalgamable gold, and a 
large proportion of fine cyanide gold, with little or no slimes, the 
method of treatment at the Crown Mines is : — 

(a.) Crushing with cyanide solution. 

(b.) Direct cyanide treatment of mixed sands and slimes by 

(c.) Plate amalgamation of free gold^ 

With an ore so exceptionally free from slimes, it seems that 
the order of treatment could be reversed with advantage, both as 
regards stamp duty and efficiency of amalgamation on the plates. 

The Crown Mines Company was the first to adopt wet-crushing 
for these gold and silver-bearing chalcedonic ores, which occurred 
in 1897, and much credit is due to Mr. F. R. W. Daw, the superin- 
tendent, for the successful inauguration of the method. 

The ore is hard and splintery, clear and pure from all impurities, 
and, unlike most of the ores from the neighbouring mines, contains 
little or no silver except what is alloyed with the gold. It is 
crushed in the Company's 60-stamp mill with cyanide solution in 
the mortars, about 2£ tons of solution being used to 1 ton of ore. 
A 2 5 -mesh screen is used, and the duty per stamp is about 2 tons 
per day. The slimes formed in crushing are said to amount to 
less than 5 per cent. 


The monthly output is about 2750 tons. For the month 
of March 2916 tons were crushed, yielding bullion valued at 
£5797. The bullion was worth a fraction under £2 per 

The cyanide plant consists of twenty-eight leaching vats, each 
22£ ft. in diameter and 4 ft. deep. Much reticence is main- 
tained as to the exact treatment, but the main features are under- 
stood to be as follows : — 

The whole of the pulp from the 60-stamps is conducted by a 
launder to one vat, and allowed to discharge into the centre 
until the vat is about half full. The pulp is then diverted to 
another vat, which is allowed to fill in the same manner. The 
mixed sands and slimes in the first vat are allowed to settle for an 
hour or two, after which the fairly clean top solution is syphoned 
off into a collecting tank, whence it is pumped up to two elevated 
tanks, from which the solution for the stamps is supplied. The 
pulp is again diverted into the first vat until the charge is com- 
plete. After settlement, the top clear solution is again drawn off. 
In this way three or four vats may be in course of filling at the 
same time. 

The settlement^ of the slimes is effected without the aid of lime, 
by allowing the solution to percolate from the bottom of the vat 
during the periods of filling and settlement. The downward 
tendency of the currents promoted by the draining from below is 
said to cause the settlement of the finest matter within a reason- 
able time. This is a point that should be noted by cyaniders 
troubled with slimy products. 

The mixed sands and slimes are treated by percolation in the 
ordinary way. The depth of each charge is about 30 inches, 
and the weight 40 tons. The strong cyanide solution is allowed 
to percolate from thirty to forty hours, while the weak cyanide 
and water washes are drawn off by the aid of an air-pump. 

The syphon used in the sand vats consists of a length of 2£ 
in. rubber hose, to one end of which is attached a short length 
of wooden batten to keep it on the surface of the solution. The 
other end is fixed, inside the vat, to a short iron pipe passing 
through the side about 18 inches above the filter-cloth. 

In the extractor-room there are five precipitation boxes of the 
ordinary pattern, divided into compartments by baffle boards ; 
and four zinc towers, consisting of wooden boxes about 6 ft. high 
and 30 in. square, set on end and connected in a series like 
charcoal towers. The solutions flow upwards through the zinc 
turnings, the overflow being conducted in a pipe to the bottom of 
the next tower, and so on to the last. 

The sands are sluiced out of the vats over a wide expanse of 



amalgamated copper-plates, which catch a certain proportion of 
the free gold. 

The actual recovery from all sources is said to vary from 84 to 
87 per cent., but the costs are not obtainable. 

G. For ores containing some easily amalgamable gold, and fine 
gold associated with pyrites and silver sulphides, the treatment 
used is : — 

(a.) Crushing with water. 

(b.) Plate amalgamation. 

(c.) Spitzlute separation of fine slimes, if necessary. 

(d.) Vanner concentration of sulphurets. 

(e.) Cyanide treatment of sands by percolation. 

(f.) Cyanide treatment of slimes by agitation and decanting. 

(g.) Cyanide treatment of concentrates by agitation. 

The practice at the Woodstock mill is a typical example of this 

The ore is a chalcedonic and finely crystalline quartz, contain- 
ing a small proportion of clayey matter and a little pyrites. It is 
stained a greyish and blackish-brown colour through the presence 
of iron and manganese oxides. • 

At the Company's 40-stamp mill the monthly output is about 
1100 tons, the stamp duty being slightly under 2 tons of 2240 
lbs. per day. For the June month, 1901, 1000 tons of ore were 
crushed for a return of £1362, which is equal to a value of 
j£l, 7s. 6'68d. per ton. The value of the bullion varies from 8s. 
to 12s. per ounce, being principally composed of silver. 

The ore is crushed with water and passed over amalgamated 
copper-plates, from the end of which the pulp is raised by a wheel 
elevator to a spitzlute. The slimes from the spitzlute are con- 
ducted to a slime tank, while the sands carrying some heavy 
slimes are passed over vanners, which collect about 1 per cent, 
of rich concentrates. 

The vanner tailings, composed principally of sands and heavy 
slimes, are led to the leaching vats, which are provided with 
automatic distributors. The construction of the distributors is of 
the simplest character, being similar to those formerly used at 
Waihi. They consist of a central wooden box, pivoted on a 
wooden pillar fixed in the centre of the vat, and from" which 
extend seven narrow wooden launders or arms of light make and 
different lengths, so as to effect an even distribution of the pulp. 
At the end of each arm there is fixed a piece of sheet zinc to 
divert the stream to one side. 

The whole of the pulp from the vanner ends is collected in one 
stream and diverted into one vat at a time until the charge is 


filled. During the filling the overflow carries the lighter slimes 
into the slime vats. 

The sands and heavy slimes are treated with cyanide by 
ordinary percolation. 

The slimes from the spitzlute, and those from the sand vats, are 
agitated with cyanide in vats provided with slowly revolving 
arms. When the gold is dissolved lime is added, and the slimes 
are allowed to settle, after which the clear solution is decanted 
off. The slimes are washed by agitating with successive portions 
of water and decanting. 

The concentrates have a value of £30 to ,£40 per ton, a large 
proportion of the value being in silver sulphide. They are 
treated by agitation with a 4 per cent, solution of cyanide for 
thirty-six hours. Two pounds of lime are added to every ton of 
concentrates. The charge weighs about 1J tons. Mr. F. Eich, 
the superintendent, who adopted the present treatment, informed 
the author that the recovery varied from 90 to 94 per cent., at a 
cost of 18s. per ton for labour and material. 

The recovery by cyanide from all sources per ton of ore milled 
is said to vary from 82 to 86 per cent., at a cost of 4s. 9d. 

D. For very slimy ores containing very little easily amalgam- 
able gold, and a large proportion of extremely fine gold besides 
the usual silver sulphides, the treatment is as follows : — 

(a.) Crushing with cyanide solution. 

(b.) Spitzlute separation of sands and slimes. 

(c.) Treatment of sands by percolation. 

(d.) Treatment of slimes by agitation and decanting. 

This method of treatment is subject to various modifications as 
regards mechanical appliances and methods of application, but 
the general principles are everywhere the same. 

The procedure at the Waitekauri 40-starap mill is as follows : — 

The ore, which contains a good deal of oxidized products, is 
crushed with cyanide solution in the mortars. From the screens 
the pulp is conducted direct to the sand vats, into which it is 
distributed by means of revolving wooden box launders actuated 
from a secondary shaft. The slimes, of which there are about 33 
per cent., are allowed to drain into a collecting vat, flowing 
through a pipe fixed in the inside of the vat. This pipe has a 
movable joint, and is raised by a screw as the pulp accumulates 
in the vat. 

The collecting vat is provided with revolving arms which keep 
the fine slimy pulp from settling. From this vat the slimes are 
pumped into the slime leaching vats, which are provided with a 
double set of slowly revolving arms, the lower ones having rakes 


on them and the upper ones loose pieces of sacking which drag 
through the pulp. In these vats the slimes are treated by agita- 
tion and decanting, lime being added with each wash to facilitate 

The sands are treated by ordinary percolation with first a 0*5 
per cent, solution of cyanide, and then the weak and water 

There are 12 sand vats, 14 slime vats, and 2 slime collecting 
vats each 2 2 '5 ft. in diameter and 4 ft. deep. The monthly 
output of the 40-stamps is about 2200 tons of 2240 lbs. The 
June monthly output was larger than usual, being 2543 tons, 
yielding 7220 ozs. of bullion valued at £6773, 0s. 6d., equal to a 
value of 1 8s. 9d. per ounce. The actual recovery is said to be 
90 per cent., at a cost of 5s. 6d. per ton. 

The exceptionally large proportion of slimes in this ore rendered 
the adoption of wet-crushing a knotty and difficult problem. The 
increased output, higher extraction, and lower costs are proofs 
enough of the success of the change from dry-crushing which was 
effected under the supervision of Mr. G. Davey, the superin- 
tendent, without hitch, or decrease in the monthly output, a 
matter of no little moment in these days of heavily capitalized 
public companies. 

Concluding Remarks. — Among the points most likely to attract 
the notice of cyaniders are the low stamp duty and the heavy 
consumption of cyanide. 

The low stamp duty of the New Zealand mills has often been a 
subject of discussion, but no satisfactory explanation has yet been 
advanced. In the opinion of the author, it is due to the circum- 
stance that the mills having been designed and erected in the first 
place for dry-crushing, the mortars are too narrow and restricted 
to give good results by wet-crushing, and until these are replaced 
by new mortars, specially designed for wet-crushing, it seems hope- 
less to look for better results. The advantages of a stamp duty of 
4 or 5 tons per day instead of 2 tons are too obvious to require 

With reference to the large consumption of cyanide, it is well 
known that silver in all its forms requires a stronger solution to 
effect its dissolution than gold ; and in the Hauraki Goldfields the 
large consumption of cyanide is due to the presence of silver 
sulphide (principally Ag 2 S), which is generally worth saving, aud 
to the circumstance that the free gold is alloyed with about one- 
third its weight of silver. ^ 

According to Eisner's equation for the dissolution of gold by 
potassium cyanide, 4 lbs. of cyanide should dissolve 100 ozs. of 
gold, but in practice it is found that it takes nearly forty times 


that quantity. To dissolve 100 ozs. of silver would require 7*5 
lbs. of cyanide, according to the equation : — 

4Ag + 8KCy + 2 + 2H 2 = 4(AgKCy 2 ) + 4KHO. 

For the dissolution of 100 ozs. of silver existing as the sub- 
sulphide (Ag 2 S), 7*01 lbs. of cyanide would be required by the 
following equation : — 

Ag 2 S + 4KCy = 2(AgKCy 2 ) + K 2 S. 

The potassium sulphide resulting from the dissolution of silver 
sulphide also tends to cause a loss of cyanide. It has been shown 
by Crosse and others that a trace of alkaline sulphide in cyanide 
solutions does not act injuriously, but the large quantity of K 2 S 
liberated in the treatment of the silver-bearing ores of the Hauraki 
Goldfields must cause the precipitation of a portion of the dissolved 
gold in the vats. Much of this precipitated gold is doubtless 
redissolved by the excess of free cyanide present in the solutions, 
but it always requires this excess to obtain adequate extractions, 
thus necessitating the use of comparatively strong solutions. 

One of the most perplexing features connected with the treat- 
ment of these ores is the constantly varying proportion of silver, 
which necessitates the use of solutions of varying strength to. 
obtain adequate extractions, thus adding another source of anxiety 
to the many worries which the use of cyanide entails on even the 
successful cyanider. 

At many of the Hauraki mines, cyanide treatment was adopted 
by the owners on the author's recommendation, but only after he 
had made a careful investigation of the constituents of the ore, 
and repeated trials on a working scale, at the Government Metal- 
lurgical Works at the Thames. In other cases, the necessary 
experimental trials were made by the author's assistants at the 
mine. In no case was the process adopted until success had been 
assured, a precaution which doubtless has been a potent factor in 
promoting the popularity of cyanide treatment in this country. 

The ores of Te Aroha and Monowai are generally very refrac- 
tory, containing free milling gold, mostly very fine, associated 
with sulphides of silver, iron, copper, lead, zinc, and often mer- 
cury. Many attempts have been made to treat them by cyanide, 
but without success, and, so far as our present knowledge goes, it 
is doubtful if they can ever be treated successfully in the raw 
state by that process. 

For the treatment of cupriferous ores and concentrates from 
the Jubilee, Sylvia, and Monowai mines, which could not be 
treated successfully by ordinary cyaniding, the author obtained 
good results by first subjecting the ore to a chloridizing roast, and 



then leaching out the copper chlorides with water. After an 
alkaline and water wash, the gold and silver contents were ex- 
tracted by cyanide by ordinary percolation. Daring the roasting 
the silver sulphides present were chloridized, the chloride being 
easily dissolved by cyanide. 

From a large parcel of Monowai ore, 92 per cent, of the gold 
and 85 per cent, of the silver were extracted, the composition of 
the ore being (F. B. Allen, M.A., B.Sc.) :— 

Insoluble gangue, 
Copper pyrites, 
Iron pyrites, 
Galena, . 
Alumina, . 
Water and loss, 



The bullion contents of this ore were : — Gold, 1 oz. 5 dwt. ; 
silver, 14 oz., per ton. 

The Moanataiari Co.'s Works, situated at Thames, are 
of recent construction, and in point of completeness and modern- 
ness of equipment, among the finest in New Zealand. The plant 
consists of a 60-head battery, two rock-breakers, grizzlies, auto- 
matic feeders, 24 vanners, 21 Berdan pans, a complete cyanide 
plant for treating vanner concentrates, 9 Cornish buddies for 
concentrating the tailings from the vanners, and all up-to-date 
appliances for assaying and retorting. The cyanide works consist 
of three steel vats, 20 ft. in diameter and 7 ft. deep, each provided 
with two bottom-discharge doors, with a capacity each of 200 
tons of concentrates ; two zinc extractors, each 15 ft. 6 in. in 
length ; and three concrete sumps, 50 x 1 1 x 6 ft. over all. 
Wet-crushing, concentration, and the cyanidation of the con- 
centrates are the interesting features of the practice at these 
works, which were designed by the author and erected under his 
supervision. The cyaniding of concentrates is comparatively new 
in New Zealand. 

Talisman Mine. — In the Ohinemuri Goldfields, the practice 
of dry-crushing and direct cyanide treatment has been superseded 
by wet-crushing, concentration, and cyanide treatment of sands, 
slimes, and in some cases the concentrates. At the Talisman 
mill, the ore was conveyed from the mine by an aerial tramway 
being dumped on to a grizzly, which passed the coarse stuff to a 
Blake Marsden crusher, whence the ore , passed to a revolving 


drying furnace. The stamps weighed 1000 lbs., and the mortars 
were provided with back and front discharge. 

The pulverized ore was elevated to an ore-bin, from which it was 
carried to the cyanide vats by a long line of revolving screw^ 
conveyors. The mechanical drier was a very efficient machine, 
drying about 12 tons of ore for every ton of firewood consumed. 

This company's dry-crushing mill has been dismantled, to give 
place to a new 50-stamp wet-crushing plant. Wire plate-amal- 
gamation, vanner concentration, separation of slimes, cyaniding of 
sands and slimes, the latter by agitation and decantation, has been 

Waihi Mine. — The ore in the upper levels of the celebrated 
Martha lode at the Waihi mine is typical of most of the ores in 
this district. It consists principally of hard, splintery, whitish- 
grey chalcedonic and crypto-crystalline quartz, often possessing 
a banded and wavy structure. It is perfectly free from all base 
metallic sulphides, and the amount of iron oxides present is so 
small that when roasted and pulverized the colour of the dust is 

The value varies from <£4 to £& per ton, the precious metals 
existing in the proportion of about 3 oz. silver to 1 oz. gold. The 
free gold is alloyed with about 35 per cent, of silver, being 
valued at about 53s. per oz. The greater proportion of the 
silver exists in the form of the bluish-grey sub-sulphide known as 
argentite. In the surface levels, thin leafy plates of gold were 
not infrequently seen adhering to the surface of large cuboidal 
masses of quartz, but in the lower levels a colour is rarely seen, 
the gold existing in an extremely fine state of subdivision. Such 
an ore is theoretically perfect for cyanide treatment, and actual 
experience has proved it to be so. By the old stamp battery, and 
copper-plate amalgamation, the recovery amounted to only some 
4 dwts. per ton, equal to about 15 per cent, of the value. By 
dry-crushing and pan-amalgamation the extraction was raised to 
60 per cent., but at present the actual extraction by the cyanide 
process amounts to over 90 per cent, of the assay value. 

The following particulars of the dry method of treatment 
formerly in use at the Waihi mills were supplied by Mr. H. P. 
Barry, the general manager, for the annual report of the New 
Zealand Mines Department for 1894 : — 

Drying the Ore. — The ore is trucked to the drying kilns which 
consist of open circular holes excavated in the solid rock, their 
dimensions being 37 ft. in depth and 20 ft. in diameter at the 
top, and tapering somewhat at the bottom. Each kiln is capable 
of holding 100 tons of ore at a charge. The lower part is lined 
with bricks, and finished off with a brick arch, having a door 


and an iron chute for discharging the dried ore into trucks, 
which have access to the kiln by means of a tunnel cut in the 

Charging the Kilns. — The kilns are charged with alternate 
layers of wood and ore, the layers of wood being about 5 ft. apart. 
When the kiln is fully charged, the wood is lighted, and, after it 
is all burned up, about half the charge, that is about 50 tons, is 
withdrawn and another 50 tons of raw ore, together with the 
necessary wood, are placed on the top. After this about 50 tons 
of ore are withdrawn every third day, while a similar quantity of 
raw ore and wood is added. 

The method of drying the ore is very expensive, as one ton of 
wood will only dry about three tons of ore. The cost of firewood 
at the Waihi big mill is 2s. per ton of ore dried, and the total cost 
of drying, including labour, is 2s. 6cL per ton. 

Crushing and Pulverizing. — From the kilns, the dried ore is 
trucked to the rock-breakers, whence it passes by gravitation to 
the self ore-feeders. The pulverizing machinery consists of a 
90-stamp battery and an Otis ball-mill, having a capacity of 
about 10 stamps. The ore is passed through a 40-mesh screen. 

Filling the Cyanide Leaching Vats. — From the screens, the dry 
dust falls into a long, narrow trough running parallel with the 
stamp-motors, along which it is conveyed to the dust-bin at one 
end of the mill by means of an Archimedean screw. From the 
dust-bin the pulverized ore is lifted by a bucket-belt elevator and 
discharged on to an 8 in. rubber-belt provided with rope edges, 
and by this conveyed to, and across, the dust-hopper, which is 
110 ft. long, running the entire length of the cyanide-plant house. 

The dust-hopper has twenty doors for discharging the dust into 
the trucks, which are run straight out over the leaching vats on 
to travellers running on rails. The travellers are provided with 
hand traversing gearing, thus enabling a truck to be tipped at 
any part of the vat. This is an important feature, as the finely 
pulverized material has a tendency to pack if moved about or 
touched in any way after being tipped into the vat. 

As a further preventative against packing, there is a small 
traveller fixed below the main traveller, provided with a platform 
at the height the ore has to be filled up to. All the trucks are 
tipped over this platform, which breaks the fall of the dust, and 
throws it in a light shower all around. 

The Cyanide Treatment. — The following particulars of the 
cyanide treatment were kindly supplied to me by Mr. E. G. Banks, 
the chemist of the cyanide operations. 

The plant consisted of thirty-eight circular leaching vats, each 
22 J ft. in diameter and 4 ft. deep, together with the necessary 


dissolving and solution vats, sumps, extractors, vacuum-cylinders, 
solution and air-pumps, etc. : — 

Filling vat, 30 tons, two men, 2£ 

Strong solution, 10 tons, 0*35% KCy leaching, . 30 

Weak solution, 7 tons, 0*1% KCy, with vacuum, 15 

First water wash, 6 tons, with vacuum, . . 24 

Second „ „ „ . 36 

Discharging vat, one man sluicing, ... 2 
Taking up and cleaning filter-bottom, . . 4£ . 

Total, . . .114 


A vacuum of 20 in. to 23 in. is maintained to obtain the above 
results. The average value of the ore was about «£4 per ton, and 
the actual extraction from 90 per cent, to 92 per cent, of the 
original value, at a cost of 7s. 6d. per ton, not including royalty. 

Dry-crushing and direct cyaniding of the pulverized ore has 
been abandoned at all the Waihi Company's mills in favour of 
wet-crushing, concentration, and subsequent treatment of sands 
and slimes. The treatment of the slimes is effected by agitation 
and tilter-pressing, combining some of the essential features of 
both South African and Western Australian practice. 

I am indebted to Mr. G. Banks, the company's metallurgist, for 
the following clear and succinct notes on Waihi cyanide practice 
at the different mills. Victoria mill. — The ore is pulverized by 
200 stamps to pass through 40-mesh wire wove steel screens, and 
is then elevated by means of plunger sand pumps to launders 
which convey the pulp to nests of spitzluten where the separa- 
tion of the sands and slimes takes place. 

Treatment of Sands. — The sands flow to the intermediate sand 
collecting vats, of which there are five built of steel, 38 ft. in dia- 
meter by 8 ft. deep, fitted with an annular launder on the outside 
top edge to convey the overflow water, at times containing a little 
slime, away to the slime thickening boxes. The sands are charged 
into the percolating vats by means of an automatic revolving dis- 
tributor, which is moved on an overhead traveller from vat to vat 
as required. 

The vats are fitted with Roche's bottom-discharge doors. The 
filter-bed is arranged by a wooden grating covered with wool-pack. 
Each vat holds from 250 to 300 tons of sands. After draining to 
get rid of surplus water, a preliminary treatment is given with 
weak cyanide solution, followed by a strong (0*35 per cent, to 
Q*45 per cent.) solution and usual washes, 


The old vat-house contains ten rectangular concrete vats, five 
on each side. Each vat is 50 ft. by 40 ft. and 4 ft. deep. 

Slime Treatment. — The slimes from the spitzluten and the over- 
flow from the sand vats are mixed with lime-water (about 2 to 
4 lbs. of lime per ton of slime) and conveyed to a nest of 36 V-shaped 
thickening boxes in which the sliinea quickly settle, and are then 
drawn off in a fairly thick state from the bottom of the boxes. 
The overflowing clear water is returned to the stamps. 

The thickened slimes are collected in six steel vats 32 ft. in dia- 
meter by 14 ft. deep. In these the slimes rapidly settle, and as 
the clear water overflows, it is stored for re-use in the mill and 

When the vat is filled with slime-pulp to within 2 or 3 ft. from 
the top, the inflowing slimes are cut off and the charge allowed to 
settle for about 24 hours. Whatever clear water may be on top 
is syphoned off, and the thickened pulp (now about 1 of slime to 
1 of water) is ready for treatment in the agitators. 

The agitators are built of steel, 20 ft. in diameter by 6 ft. deep, 
and hold about 25 tons of slime (dry weight), together with about 
40 tons of cyanide solution. The strength of the solution is about 
O'l per cent. The pulp is agitated by paddles, secured to a vertical 
shaft, actuated by overhead worm-gearing. The speed is 8 
revolutions per minute. 

After 24 hours' agitation the pulp is run to a pressure tank, 
and from thence is forced by compressed air into Johnson 6-ton 
filter-presses, where the gold bearing cyanide solution is extracted. 

The zinc method of precipitation is used and does very good 
work, even on very dilute solutions. 

Vanners are being erected to concentrate the more heavily 
mineralized ore. The tailings from the vanners will be mixed 
with the lightly mineralized ore-pulp, and pass through the course 
of treatment just described. 

An extraction of 85 per cent, to 90 per cent, of the ore value 
can be recovered by this process, but no particulars of the actual 
recovery or costs are yet available. 

In the month of January of this year, the Waihi Company 
crushed 12,968 tons of ore, which yielded by cyanide treatment 
bullion valued at £29,055. 

Waihi Union Mill and Cyanide Plant. — This mill consists 
of 40-stamps for wet-crushing and plate-amalgamation ; sixteen 
circular steel vats for sand treatment (each vat having a capacity 
of 22 tons of sand) ; three slime-collecting vats, 32 ft. in diameter 
by 14 ft. deep; two agitators, 20 ft. in diameter by 6 ft. deep; 
and one Johnson filter-press, with a capacity of 6 tons of slime- 
dry weight, per charge. In addition there are the usual ore^ 

i 1 




I . 



J — 10 — 




Scale of Feet 



11 1 " 1 """ 1 



breakers, spitzluten, pumps, air compressors, etc. The screens 
used are 40-mesh wire-wove steel, and the pulp after passing the 
amalgamated tables is elevated by a wheel-elevator to a height of 
25 ft. and is here classified into sands and slimes by four spitzluten. 

Sand Treatment — The sands flow to the vats and are distributed 
by means of an automatic revolving distributor. A weak sump- 
wash, containing about 0*07 per cent, to O'l per cent, of cyanide, 
is run on, and this is followed by a solution 0*4 to 0*6 strong. 
The total time of treatment is from five to six days. The tailings 
are sluiced away through a bottom central discharge-door. 

Slime Treatment. — The slimes, after the addition of lime-water 
(about 3 lbs. of lime per ton of ore), flow to the large collecting 
vats and are there thickened by natural settlement to about one 
part of slime to one of water. In these vats, a syphon provided 
with a ball-and-socket joint drains off the top clear water 
automatically. The clear water overflowing from the settling vats 
is sent to a reservoir to be pumped up for re-use in the mortar- 
boxes. This course is found to be necessary owing to the fact 
that even when the water appears to be perfectly clear it contains 
from £ grain to 2£ grains of gold per ton, and as several tons of 
water are used per ton of ore, it is evident the loss would be from 
6d. or 8d. up to two or three shillings per ton of ore, if the water 
were allowed to run to waste. 

From the collecting or settling vats, the thickened slimes are 
drawn off to the agitators, situated at a lower level, in charges 
equal to 20 to 25 tons of dry slime, and are agitated for 24 hours 
with 1 J tons of 0'1 per cent, cyanide solution to 1 ton of dry slime. 

Waihi Mill. — This is the oldest of the Waihi Company's 
mills. Here the ore for some years was dry-crushed and pan- 
amalgamated by the Washoe process ; and afterwards dry-crushed, 
and the dry pulp cyanided by the direct process in the manner 
described by Mr. Barry in the preceding pages. Of the present 
treatment Mr. Banks writes in April of 1903 as follows : — 

" At the old 90-stamp mill dry-crushing was stopped at the 
end of 1902 and alterations necessary for wet- crushing rapidly 
completed. By the middle of January 1 903, the mill was restarted. 
The ore (about 220 tons per day) is stamped through 40-mesh 
wire- wove steel screens ; passed over amalgamated tables, of which 
there are fifteen, each 12 ft. by 6 ft. ; and then over thirty union 
vanners which extract about 2£ to 4 per cent, of concentrates. 
After passing the vanners, the pulp is elevated by wheel elevators 
to a series of sand separating boxes. Here the sands and slimes 
are separated, care being taken to keep the sand as free from 
slime as possible. 

f< Treatment of the Sand. — This is effected in circular vats 4J ft 


deep and capable of dealing with 40 tons of sand per charge. The 
vat is first filled with water and the sand is then run in through 
an automatic revolving distributor. The overflow water (carrying 
a little fine sand and slime) runs into an annular launder and is 
conveyed back to the elevating wheels and joins the pulp flowing 
to the separating boxes. When sufficient sand has been run into 
the vat a 2 in. draw -off pipe, passing through the side of the vat, 
is lowered and the water drawn off. This water also contains a 
little slime and flows to the elevating wheels. 

" After the charge has drained for several hours, a weak sump 
wash (testing about 0*08 per cent. KCy) is run on, and this is 
followed by a strong solution testing 0*5 per cent. KCy ; when 
this solution has reached the bottom of the charge, percolation is 
stopped and the solution circulated by means of an air-lift inside 
the vat. After five days the strong solution is drawn off and the 
charge washed with weak solution and water. The tailings are 
then sluiced away. 

" Treatment of the Slime. — After separation from the sand, the 
slime pulp (about 20 of water to 1 of slime) is mixed with lime- 
water and conducted over a nest of spitzkasten, thickening the 
slime-pulp to about 10 per cent, of slime, the balance of the water 
passing over in a clear state to be re-used in the mortars. 

" The slime-pulp is now collected in one of two collecting vats, 
which are 14 ft. deep by 32 ft. in diameter, and fitted with an 
overflow launder. The pulp is run to the centre of the vat and 
discharged through a vertical box chute extending several feet below 
the top of the vat, so as to cause as little disturbance as possible. 
The pulp runs in until the slimes show at the top, the clear water 
overflowing to a reservoir for re-use. The collecting vats thicken 
the pulp to about 2 J of water to 1 of slime. 

" The thickened slime pulp is now still further separated from 
the contained water by means of filter-presses, of which two 
are required for * drying ' the slimes, each dealing with about 40 
tons (dry weight) of slime per day. 

" The slime-cakes, now containing only 25 per cent, to 35 per 
cent, of moisture, drop from the presses on to a screen-conveyor 
which discharges into a disintegrator. This disintegrator is 7 ft. 
in diameter by 14 ft. deep and fitted with three sets of revolving 
arms bolted to a centre shaft driven by overhead gearing at the 
rate of 20 revolutions per minute. Cyanide solution (0*12 per 
cent.) flows in at the bottom of the disintegrator in such proportion 
that the overflowing pulp contains 1 of slime to 1£ of solution. 
This pulp flows through a series of four agitators each 14 ft. deep 
by 20 ft. in diameter, fitted with stirring gear making three 
revolutions per minute. 


" The stirring gear consists of two arms secured one foot from 
the bottom of the vat to a central shaft driven by overhead 
worm-gearing. Effective agitation is produced by blowing com- 
pressed air through a number of pipes dipping down close to the 
stirring arms. The revolving arms serve to bring the pulp in 
contact with the escaping air, at which point fairly violent agita- 
tion is produced. 

"The pulp enters No. 1 agitator at the top, flows from the 
bottom of No. 1 through a connecting pipe terminating half way 
up the side of No. 2. Similarly it passes from No. 2 to No. 3, 
and from No. 3 to No. 4. From No. 4, the pulp flows to a larger 
agitator, No. 5, which is 32 ft. in diameter by 14 ft. deep, and 
also fitted with stirring gear and air-agitation. The compressed 
air in the pressure tank after filling a filter-press is utilized by 
being blown through the agitators. 

" The pulp is drawn off from No. 5 agitator, from time to time 
as required, to fill one of the four filter-presses used to complete 
the treatment. A press charge contains about 5 J tons (dry weight) 
of slime. The cakes are 3 in. thick. 

Time required to fill press at 60 lbs. pressure, . 1 

to wash at 75 lbs. pressure, . . 3 
to open, discharge, and close press, . 1 J 

Total, . 5£ 

" From five to six hours can be taken as the time required to 
treat a press charge of Waihi slime. Of course, the more fine 
sand contained in the slime, the quicker a press will fill and 
wash. About 25 dry tons of Waihi slime can be taken as the 
average amount treated per press per day. The labour required 
is one man to each press each shift of eight hours. The cost 
for cloths, rubber-rings, etc., is from 2d. to 4d. per ton of slime 

"Extraction. — By amalgamation and concentration from 40 
per cent, to 50 per cent, of the value is recovered, each process 
extracting about half that percentage. The sand treatment 
extracts about 80 per cent, of the gold and 60 to 70 per cent, 
of the silver contained in the sands. There is obtained by 
the slime treatment a recovery of 95 to 96 per cent, of the 
gold, and about 70 per cent, of the silver. The total extraction 
is about 89 per cent, to 90 per cent, of the gold and 75 per 
cent, of the silver. 

" The consumption of cyanide per ton of ore varies from 2 J lbs. 
on lightly mineralized ore, to 4 lbs. on heavy sulphide ore. 

142 wis CTAHIM MtOCftSS. 

"The cost of thd siime treatment is not given, but it will 
probably work out to about 6s. 6d. per ton of slimes treated. 

" The output of the Waihi Company is about 18,000 long tons 
every four weeks of twenty-four working days, for a return 
varying from £50,000 to £52,000. The ore is pulverized in 
the company's mills, comprising 320 stamps, and then subjected 
to cyanide treatment by the various processes described above. 
The cyanide plants have a capacity of about 1000 tons per 

From the foregoing description it will be seen that the slime 
process in use at the Waihi and Waihi Union mills is practically 
the same as that in use at Kalgoorlie in Western Australia. On 
the other hand, the process at the Waihi Company's Waikino mills 
is almost identical with that in use at Glencairn Main Reef, 
Johannesburg, differing only in the final stages in the adoption 
of filter-presses to separate solutions from the slimes instead of 
decantation which is almost universal in South Africa. 

Waihi Tailings. — The treatment of the Waihi Company's 
tailings is very instructive. The tailings are the residues result- 
ing from the pan-amalgamation of the dry -crushed ore before the 
introduction of the cyanide process. The ore was crushed through 
a 60-mesh screen, and pan-amalgamated in charges. The residues 
were discharged from the settlers into large dams, where they 
were allowed to settle. They mostly consisted of fine sands and a 
good deal of slimes. They contained no base metallic impurities, 
and the gold existed principally in the form of amalgam. 

Some 25,000 tons of these tailings were successfully treated by 
the Cassel Gold Extracting Company, whose works have recently 
been acquired by the Waihi Gold and Silver Mining Company, 
who treated the remainder of the tailings on their own account. 
The plant consists of eight leaching vats, each 22£ ft. in diameter 
and 4 ft. deep, together with all the necessary appliances. 

The details of the cyanide treatment adopted for the treatment 
of these tailings are given below in tabulated form : — 

Cyanide Treatment op Waihi Tailings. 

Filling leaching vat, 30 tons, three men, ... 8 
Preliminary lime or water- wash, 6 tons, with vacuum, 6 

Leaching — 

. Strong solution, 8 tons, 0*6% KCy, . . . .30 
Weak solution, 4 tons, 0*2% KCy (from strong sump), 12 


Washing, using Vacuum — 

First weak cyanide wash (from weak sump), 4 tons, 

Second ,, ,, „ 



> • 














Fourth, water wash, 4 tons, 
Discharging vat, one man sluicing, 

Total, . . . .108 

Remarks. — The tailings were generally very clean, and the pre- 
liminary lime or water wash was not always applied. The strong 
solution was allowed to stand in contact with the tailings about 
four hours before the percolation was commenced. The average 
value of the tailings was about 24s. per ton, and the actual 
extraction about 75 per cent., at the cost of 8s. per ton. 

Try Fluke Mine. — At the works of this Company, at Kuaotunu, 
the ore was wet-crushed through a 40-mesh screen, and passed 
over amalgamated copper-plates. The tailings were run directly 
from the plates into settling-pits. When the pits were full, the 
slimes were removed from the lower end, and spread out to dry in 
the sun. When dry, they were broken up, and then were filled 
into] the leaching vats together with sand, in the proportion of 
one truck of slimes to two trucks of sand. 

The sand and slimes were thoroughly mixed in the vat before 
the solution was put on. The Company's metallurgist informs the 
author that the average time of treatment was as follows : — 


Filling vats, 20 tons, 6 to 8 

Strong solution, 5 tons, 0*6% KCy, standing 

in contact with tailings, . . . . 8 to 1 2 

Percolating, . . . . . . 24 to 30 

Weak solution, 5 tons, 0*2% KCy, . . 5 to 6 
Weak cyanide washes, five of 5 tons each, 

0-1% of KCy, 25 to 30 

Totals, . . . . 68 to 86 

The ore consisted of grey and yellowish-brown quartz, sometimes 
containing a considerable proportion of iron and manganese 
oxides, the latter generally predominating. The greater part of 
the gold was excessively fine, being locally known as " float gold." 

The average value of the tailings was about 20s. per ton ; and 
the actual extraction by cyanide about 75 per cent., at a cost of 
7s. 6d. per ton. When the tailings were of higher value than 
usual, they were turned over in the vat after the last washing and 


washed again. In this case it was found that the extra extraction 
more than paid for the extra labour. 

At the Kapai- Vermont mine, which adjoins the Try Fluke, the 
same ore was dry-crushed in a ball-mill, and then subjected to 
direct cyanide treatment with the most satisfactory results, the 
actual extraction generally exceeding 85 per cent, of the assay 
value. In this mine, shoots of very rich ore have been frequently 
met with, containing a considerable proportion of comparatively 
coarse gold. With such ores the strong cyanide solution was 
circulated through the leaching vats until an adequate extraction 
was obtained. 

Waitekauri Mine. — At the Golden Cross section of the 
Waitekauri Gold Mining Company's Special Claim, near Waihi, 
the ore, before the introduction of wet-crushing at the new mill, 
was dry-crushed with stamps to pass through a 40-mesh screen, 
and treated directly with cyanide. As a small proportion of the 
gold was coarse, the tailings were passed over amalgamated 
copper-plates, 30 ft. long and 3 ft. wide, set with a fall of 1 in 12. 

The details of the cyanide treatment of this ore may prove 
useful, and are given in the following tabulated statement : — 

Filling vats, three men, 22 tons, .... 3^ 
Strong solution, 9 tons, 45% KCy, ... 48 
Weak solution, 9 tons, 0'2% „ > . . .18 
First weak cyanide wash, 5*5 tons, 0*05 to 0*15%, 18 
Second „ „ 5*5 „ „ .18 

Third „ „ 5'5 „ „ .18 

Fourth, water wash, 3*0 „ . . .18 

Discharging vat, one man sluicing, . . 14 J 

Total, . . . . .156 

The average value of the ore treated during 1894 was £4, 15s. 
per ton; and the actual extraction varied from 91 per cent, to 93 
per cent., at a cost of 8s. 6d. per ton. 

Treatment of Sulpho-Telluride Ores. 

At Kalgoorlie, the chief mining centre in the State, the gold 
occurs in the oxidized surface ores in a free state, and in the 
unoxidized ores in combination with tellurides and sulphides. 
In all cases it exists in an extremely fine condition in talcose 


calcareous ores that possess a great propensity to form slimes. 
The successful treatment of these ores at first presented a 
difficult problem to metallurgists, and after repeated failures 
along old lines, a process was developed that in some respects 
possesses some peculiar features. And the early difficulties 
were not confined to the mere treatment of the rebellious ore. 
The climate was tropical and dry, water scarce and often brackish, 
and the distance from the seaboard great, and over arid plains. 
But these difficulties have now been overcome. A State railway 
connects the distant mining centres with the capital, and a State 
water supply, the greatest .undertaking of the kind in the world, 
provides an abundance of pure water, carried in a pipe-line for a 
distance exceeding 300 miles. 

The telluride ores are very brittle, and consequently a larger 
proportion of the gold goes into the slimes than into the sands. 
This circumstance and the slimy character of the ore led to the 
adoption of the cyanide process and filter-press treatment of the 
slimes. The object at many of the mills is to slime as high a 
proportion of the ore as possible. 

The Diehl and Riecken processes have been installed at several 
mines with, it is said, satisfactory results. A detailed description 
of them will be found in Chapter XIII. 

In some of the mills the slimes are drained or partially dried in 
presses, disintegrated and agitated with cyanide solution, and 
afterwards pressed and washed in presses. In other mills, the 
dissolution of the gold is effected directly in the presses. 

I am indebted to Mr. F. B. Allen, M.A., B.Sc, Director of the 
Kalgoorlie School of Mines, for the following interesting details of 
the general methods of treatment adopted at that mining 
centre : — 

General. — The essential features of the treatment are dry- 
crushing, roasting or not according to circumstances, fine sliming, 
leaching with cyanide, and filter-pressing. 

The roasting is generally effected in Edwards or Brown 
straight-line furnaces. . The former roast from 14 to 16 tons per 
day, at a cost of 4s. to 4s. 6d. per ton ; and Brown, 30 tons per 
day to 0*1 per cent, sulphur as sulphide for 7s. to 9s. per ton. 

Ore containing over 3 per cent, of moisture is dried before dry- 
crushing in ball-mills. No 5 Krupp, running at 25 revolutions 
per minute with 15 H.P., crushes 25 tons through a 40-mesh 
screen for Is. 2d. per ton, and of this from 65 to 75 per cent, will 
pass a 100-mesh sieve. The Griffin mill, when crushing to a 15- 
mesh, will form a product of which 75 per cent, will pass through 
a 100-mesh sieve, at an approximate cost of 2s. per ton. 

Great Boulder Proprietary.— The sulphide ore is partly 



crushed wet with stamps and amalgamated in Wheeler pans, and 
partly dry-crushed in Griffin mills, roasted, fed into mixers which 
supply Wheeler pans with the pulp, which is further ground and 
amalgamated with mercury without the use of copper-plates. 

Cyanide Process. — The pulp from the continuous overflow of the 
pans is led into settlers, first passing, in the case of sulphide ore, 
over canvas tables to eliminate concentrates. About one-third of 
the gold contents are recovered by amalgamation. The furnaces * 
used are Edwards (16 tons per day), fed with producer gas, and 
they roast the ore down to 0*11 per cent, sulphur as sulphide, 
and 2*0 per cent, sulphur as sulphate. Push conveyors are used 
throughout, and samples are taken automatically as the ore is 
discharged from the elevators. 

The fine slimes from the settler go to a settling tank, whence 
they are lifted up and passed through spitzluten, the heavier 
particles being returned to the pans. The slimes, 1 to 1, are 
agitated and passed to montejus or pressure-tanks, and then to 

The four Dehne presses hold 4 tons each; the five Martin 
presses 3J tons, forming 3 in. cakes. They are worked by 
hydraulic pressure. 

The solutions are clarified by Excelsior presses, passed through 
three zinc boxes, and the gold-slimes treated with sulphuric acid 
and melted in tilting furnaces. 

The residues are dumped by a Ledger wood hoist on to a 60 ft. 

KalgOOrlie Mine. — Part of the ore is dried in a White-Howell 
drier, the drier ore passing direct to large 200-ton storage bins. 

The ore is automatically fed into six No. 5 Krupp ball-mills, 
crushed through a 35-mesh screen, so that about half will pass 
through a 120-mesh, and then led into a 400-ton bin, whence by 
roll feeders it is fed into nine Edwards roasters, each treating 
15 tons per 24 hours. 

The roasted ore falls on to a push-conveyor and is carried to a 
bucket-elevator, which lifts it to a mixer. The pulp from this is 
separated by a series of conical spitzkasten into slimes, which pass 
on through several pyramidal spitzkasten to have the surplus 
water removed ; while the sands and concentrates run over 
copper-plates 10 ft. long and Halley tables, from which the con- 
centrates are ground in Wheeler pans. The sands are sent to 
three 100-ton vats, drained of water, and bottom-discharged into 
steel vats below. Ordinary cyanide treatment occupies three 
weeks, and the cyanide solutions as they come off are returned to 

* Three-hearth Merton furnaces are now being used. 


the top of the sands by an air-lift, which keeps the solutions 
constantly circulating. 

The slimes from the spitzkasten thickened up to 1 in 1 flow to 
a set of five pressure tanks 5| x 13 J ft., where they are agitated 
with compressed air and cyanide solution for four hours. One 
tank is being filled while another is being emptied and the others 
working, the air passing from one to the other. The agitated 
pulp is then filter-pressed. Cyclone settling cones are used for 
settling the dust, as well as an air-lift for the pulp. 

Golden Horseshoe. — The oxidized ore is screened through 
2 in. grizzlies, the coarser lumps passed through No. 3 Gates 
crushers, and the broken ore stored in a 200-ton bin, whence it is 
trucked to Challenge ore feeders supplying a 50-stamp battery of 
1000-lb. stamps. The ore is wet-crushed through a 24-mesh 
woven wire-screen at the rate of 4 or 5 tons per 24 hours, 
the greater part passing a 100-mesh screen. Free gold is 
amalgamated both inside the boxes and outside on copper-plates, 
which are followed by concentrators of the Wilfley type. The 
coarse sands are ground and amalgamated in a series of pans 
followed by settlers and spitzkasten. The fine sands thus 
obtained are lifted by a 42 ft. tailings-wheel to vats fitted with 
Butters distributors. The slimes pass over 30 ft. of canvas 
tables for the further elimination of fire concentrates. Double 
treatment of the sands, occupying nine or ten days, is practised. 
The fine concentrates, etc., are collected and sent to the smelter. 
The slimes are pumped direct, without agitators or montejus, 
to nine Dehne presses, each carrying fifty 3 in. cakes, pressed, 
washed with cyanide solution and water, and discharged. 
Excelsior presses are used for clarifying the gold solutions, which 
are then passed to four zinc extractor boxes. 

Sulphuric acid treatment with filter- pressing is adopted for 
treatment of the zinc slimes, which are finally melted in a tilting 

Stdphide Ore. — The new battery is of 50 stamps, each 1250 lbs. 
The ore, after passing the Gates crusher, is brought by a Robins 
belt to Challenge feeders and crushed wet through a 30-mesh 
screen, passed over copper-plates, and lifted by a tailings-wheel to 
a series of grading boxes supplying Wilfley tables. The slimes 
which overflow from the graders, together with that from the 
Wilfleys, are lifted again to three hydraulic classifiers, the slimes 
passing to settling tanks and the sands to five flint-mills which 
crush them to 220-mesh, and then are again elevated to the 
graders and finally passed to the settling vats. The thickened 
slime produced in these vats is run into steel agitators, agitated 
with cyanide for 24 hours, pressed in six 5-ton presses, and the 


waste material dropped on to Robins belt conveyors for removal. 
Rich sulphide ore is sent to the smelters. 

Great Boulder Main Reef. — The Sulphide ore is treated 
without drying by breaking in a No. 3 Gates crusher, delivering it 
by a Challenge feeder on to a Robins belt, and then to two Krupp 
ball-mills, where it is crushed to 30 -mesh, a large portion, especially 
when the ore is schistose, passing 100-mesh. 

The crushed ore is roasted in a Richards shaft-furnace 65 ft. 
high, with eleven floors, and rabbled by hand, especially on the 
lower floors. In this furnace 35 tons per day are treated. There 
are also three Edwards furnaces, each of a capacity of 12 or 15 
tons, which are bricked up to a constant angle. 

The hot roasted ore falls into a launder carrying dilute cyanide 
solution, and is raised by a tailings-wheel to a spitzkasten. The 
sands are ground fine in Wheeler pans containing mercury, but no 
plates, and with a continuous overflow into another spitzkasten, 
from which the sands again pass to the pans, while the slimes join 
the first slimes, which are pumped into agitation vats. These vats 
are 21 ft. in diameter and 6 ft. deep. The vats are allowed to fill 
in 7 hours, the solution gaining in cyanide during the last foot of 
filling from cake cyanide then added. 

After 10 hours' agitation the pulp is discharged or run off into 
montejus and filter-pressed in 3 in. cakes (Dehne). 

The solutions are clarified by passing through a small press, and 
led into zinc boxes. The gold slimes are washed, pressed, treated 
with sulphuric acid, dried in a large iron muffle, and smelted in 
Cornish furnaces. 

Dehne : filling in, . . . . £ 

Leaching and washing . . . 1£ 

Discharging, \ 

Four tons in, 2| 

Lake View. — Sulphides, roasted. — From Gates No. 5 crusher 
the ore passes to 400-ton ore-bin, and thence by aerial tram to 
ball-mill bin. There are four Krupp mills with a capacity of 130 
tous per day, 65 per cent, of which will pass a 150- mesh screen. 

Ore is led by conveyors to four Brown straight-line furnaces 180 
ft. x 10 ft., roasting 30 tons per day down to 0'2 per cent, of 
sulphur as sulphides. The roasted ore is elevated into 50- ton 
agitators into which 15 per cent. KCy is running, and the 
resulting sands are transferred into ten leaching vats, while the 
slimes pass on to three agitators for further agitation, before being 
passed through montejus and Dehne presses. Then follows zinc 
precipitation and sulphuric acid treatment. 


Sulphides — DieM Process.— The ore is wet-crushed, amalgamated 
on plates. The sandy concentrates from Wilfley tables are roasted 
in Edwards furnace, restamped, and ground in grit-mills down to a 
fine pulp. The pulp, after having been thickened by passing through 
spitzkasten,is treated in closed steel vats by agitation for 20 hours with 
bromo-cyanogen. Then follows filter-pressing and zinc precipit ation. 

The cost of treatment by Diehl process in 1901 amounted to 30s. 
per ton, including royalty Is. 9d., water 4s. 3d., bromo-cyanogen 
4s. 6d., cyanide 3s. 6d. 

South Kalgoorlie. — Oxidized Ore. — The sequence of opera- 
tions is as follows: — Gates crusher, Cornish rolls, Griffin mills 
(wet), copper-plates, grinding pans, canvas tables, spitzkasten, 
concentration of slimes, slimes agitation, Dehne filter-pressing. 

Sulphide Ore. — .Griffin mills (dry) ; two Brown straight-line 
furnaces, 150 ft. x 10 it., with an 80- ton capacity. The roasted 
ore is elevated and mixed .with cyanide solution, 1 to 1, and 
then treated in vats by the Riecken process. The vats are four 
in number, 11 ft. deep and about 13 f t. x 8 ft., with sloping sides 
and rounded bottom, holding about 18 tons of ore. KCy, 0*05 
per cent. Electricity — amperage, 200 ; voltage, 2 to 3. 

The current is supplied through iron anodes between the 
paddles of the agitator, which are kept moving at about 12 
revolutions per minute for 18 hours. The current passes off by the 
copper-plates suspended along the sides, which are kept bright by 
a constant stream of mercury fed from pipes just above the level 
of the pulp, and moving backwards and forwards across the plates. 
The mercury is lifted by an air-lift after being drawn off below, 
for continuous circulation. The copper-plates are easily removable 
by tackle, and with the mercury from the well yield amalgam. 

The pulp is drawn off by a bottom valve, and led into a 
storage vat 12 ft. x 6 ft., and kept agitated. Thence it is drawn 
off as required through montejus and filter-pressed. 

The raw ore has the following composition : — 


• • • 


per cent. 


• • • 



Ferric oxide, 

• • • 




Iron (in pyrites), 

• • • 



Sulphur (in pyrites), 

• • • 



Lime, .... 

• • • 



Carbon dioxide, . 

• • • 




• • • 




• • • 



Combined water, alkalies, a 

,nd tellurium, 





After roasting, analysis shows sulphur as sulphates, 2*82 per 
cent. ; sulphur as sulphide, 0*32 ; sulphur volatilised, 0'46. Total 
sulphur, 3*60 per cent. 

In a recent report on the Riecken process at the South Kalgoorlie 
mine, Mr. R. Hamilton gives the costs as follows : — 

8, d. 

1. Coarse crushing and transport to mill, . 2 7 per ton. 

2. Fine pulverising, including proportion of 

general charges and power, . .53 

3. Roasting and conveying ore, including pro- 

portion of general charges and power, . 7 Of 

4. Agitation with cyanide, and ejectrical pre- 

cipitation treatment, . . . . 8 5| 

5. Filter-pressing and washing, . . .63 



Total charges, . JE1 9 7J „ 

The average value of the ore for January, February, and March 
was £3, 17s. 6d. per ton, and the value of residues 2*125 dwt. 
of gold, equal to a value of 8s. lOJd. per ton, representing an 
extraction of 88 '4 per cent. 

Costs, February 1903. 

Ivanhoe Miiue. 

[y lowest costs on the field : — 

8, d. 
13 10-73 
11 9-65 


Reduction, .... 

Development, .... 
Capital, ..... 
General, ..... 

7 11-85 

2 3 03 

3 0-55 

38 11-81 per ton. 
11-098 tons for 10,318 ozs. of gold, valued at £43,800. 

Lake View Mine. 

s. d. 

Stoping, .... 9 5-6 

Diehl process, . . . 17 11 

General, . . . . 3 6*3 

30 10*9 working costs. 

Mine development, . . . 1 3*9 approx. 
Additions to plant, . . . 4*9 „ 

Kalgoorlie Mine, 46s. per ton (higher than usual). 

application of the process in different countries. 151 
New South Wales Filter-Press Practice. 

The following interesting details are extracted from a paper 
recently written by Mr. I. W. Rock.* The plant to which these 
details refer is one capable of dealing with 200 tons of dry slimes 
per day. A summary of the process is as follows : — 

1. The dry slimes are discharged into a mixer, in which they 
are ground up, while cyanide solution is added, the mixture out- 
flowing into a storage tank, whence it is elevated into agitator 
vats by means of a centrifugal pump. 2. Mixture agitated for a 
period ascertained by experiment, by some mechanical means, 
thus ensuring intimate contact between the metallic and chemical 
particles. 3. Contents run iuto montejus, which consist of cylin- 
drical vessels provided with inlet and outlet valves, and also 
connections to the receiver of an air compressor. 4. Forcing the 
mixture by means of compressed air from the montejus into filter- 
presses, the gold-bearing solution escaping through the filter-cloths 
of the presses, while the solid material remains in the frames, 
forming large thin cakes. 5. A further extraction from these 
cakes by "washing" or forcing, at a high pressure, a weaker 
solution of cyanide through them, and, if found necessary, a 
second washing with water only. 6. Getting rid of the exhausted 
slimes by opening the presses and discharging the cakes into 
trucks, which convey them to a dump. 7. Clarifying the ex- 
tracted gold solutions flowing from the presses, by running them 
through sand traps and thence into tanks, pumping them into 
elevated tanks, in order to obtain regular pressure, and passing 
them through a finer class of filter-presses, in order to retain any 
impalpable solid matter which would foul the zinc extractor boxes. 
8. Passing the outflow from the presses into the extractor boxes 
and treating therein in the usual manner, the outflow being of 
course returned to sumps for strengthening and re-use. 

Mixers. — These are made as a standard article by the 
engineering trade, and consist of a steel cylinder 6 feet in dia- 
meter by 5 feet deep, lined with an inserted cone, which has 
openings in it near the top of the outside shell, and at the 
bottom a four-bladed propeller on a vertical shaft, with mitre 
wheels and driving gear. The propeller not only cuts up the 
slimes and mixes them with the solution, but it drives the mixture 
upwards through the openings and up to the outlet, whence it 
overflows into the storage tank, on its way to the agitators. 

Agitators. — These consist of steel vats 16 feet high and 8 feet 
diameter, open at the top, and with the lower portions tapered to 
openings which connect with both outlet and circulating pipes. 

* Rock, The Australian Mining Standard, Dec. 12, 1901. 


To each of these agitators there is connected a centrifugal pump, 
whose sole duty is to draw the contents of the vat from the 
bottom and redeliver it into the top. This mode of mixing the 
particles and bringing them into intimate* contact has been found 
highly efficient and economical in regard to time as compared 
With arms or paddles adopted in horizontal agitators previously 

Centrifugal Pumps. — These circulating pumps, and the one for 
elevating the mixture into the agitators, are the only ones used 
for dealing with gritty material in the process. They have the 
usual water-pressure chambers in connection with the glands, but 
as water would dilute the mixture and upset the proportion, they 
are fed with cyanide solution under pressure from an elevated 
tank, supplied by a small pump. 

Montejus. — These are each large enough to receive the contents 
of an agitator, made after the style of a compressed-air receiver, 
buried vertically in the ground. Inlet and outlet valves, of 
chemical and grit-proof make, are fitted on them ; compressed-air 
supply and exhaust, test cock and pressure gauge, are fitted on 
each. In addition, there is a small air-agitation connection, so 
that any mixture remaining for a time in one is kept alive and 
the deposition of solid matter prevented. 

Filter-Presses. — These are generally made by or after the 
pattern of Dehne, the German engineer, the usual size having 
fifty 40-inch x 40-inch x 2-inch frames, the total contents being 
75 cubic feet. Jn addition to the fittings supplied by makers, 
there are several others which are necessary and have to be added, 
such as drip trays, slimes and leak launders, and several other 
minor accessories. Of course the discharge snoots for delivering 
the cakes of spent slimes, as they fall out of the frames into the 
trucks beneath, form part of the setting of the presses in the 
building prepared for them. 

Wash Solution Pump. — This may be of any description of high- 
pressure pump with cast-iron fittings. The pressure for filling 
the presses from the montejus may be reckoned at up to 80 lbs., 
that of the wash- water at 100 lbs. to the square inch. It may be 
noted here that both after filling the presses and after washing 
the contents, compressed air is admitted for a short period to 
drive out any moisture. 

Clarifying Presses. — These are also filter-presses, but of a 
much lighter description, the frames being usually made of wood. 
They are supposed to be self-cleansing, by reversal of the flow 
through them, and are so with some materials, but other slimes 
are so gluey that it is found necessary to have a spare set of 
filter-cloth frames,, which can be substituted for the foul ones in 


a few minutes, the latter being taken away and scrubbed. These 
presses are very delicate, and the pulsation of a pump destroys 
their efficiency. It is therefore necessary to pump the gold 
solution into a tank about 18 feet high, whence it flows quietly 
through the presses, and thence to the extractor boxes. 

The items of storage vats, sumps, etc., need not be particularly 
specified, and it need hardly be added that the adjuncts of suffi- 
cient steam power, compressed air, and supply of water are 

Practical Work. — If a plant as above described is doing its 
ordinary work, the routine would be as further described. . Side- 
tipping trucks are delivered at regular intervals alongside the 
mixer, and the contents dipped bodily out, the fall being broken 
by a shoot, sufficiently flat to require the attendant to use a 
shovel and in some degree regulate the feed. He also, through 
practice, knows how much cyanide solution to run in from a 
supply cock, so that a fairly constant flow of mixture escapes into 
the storage tank. 

When one of the agitators is empty, a signal is given to refill, 
and the elevated centrifugal does this in a few minutes from the 
storage tank, the circulating pump running all the time. After 
sufficient agitation, the man in charge of the machinery room 
opens the inlet valve and fills one of the montejus, the air escape 
cock being open. When the charge is all run in he closes the 
inlet valve and opens the little agitation cock until he receives a 
signal that a press is to be filled. When he gets that he closes 
the above, opens the outlet valve, and then the compressed air 
cock ; the pressure gauge rises, and in a few minutes half the 
contents of the montejus has been transferred to the press in the 
room above. In this room the signal to fill was given when any 
one of the presses had been emptied and reclosed, ready for 
filling. All the outlet cocks on the frames are open, and the 
gold solution at once flows from them until the press is full of 
solid matter. Notice is then given to close the valves previously 
open and start the wash water pump; the taps are closed, the 
wash water outlet opened, the flow therefrom directed for a little 
time into the same launder as the gold solution ran into, and 
thereafter into a main leading to a sump for re-use. Pumping is 
then stopped, a little air blown through to complete the process, 
and the press unscrewed, the drip trays being previously removed. 
One after another the dummies, as the slimes-bearing frames are 
called, are pulled forward, the cakes either fall or are pushed out 
into the trucks below, each frame is scraped clean with a scalpel, 
and the press tightened up again for a fresh charge. 

The treatment of the gold solution is pretty well automatic ; its 


flow through the clarifying presses has been already described, 
and its final delivery into the zinc boxes of the extractor-house. 
It will be understood that the diagram in no way indicates any 
arrangement of the plant. Accessibility and ample space are 
particularly necessary, and some details of the plant which are 
purchased as articles ready for use from the makers require con-, 
siderable alteration to make them suit the above requirements. 

Other States in Australia. — In Victoria, South Australia, 
and Queensland, the cyanide process is employed almost exclus- 
ively for the treatment of sands and old accumulations of 
tailings. Nowhere has any distinctive feature been developed 
except in Victoria, where charcoal is much used for precipitation 
purposes instead of zinc. This is really a survival and adaptation 
of the charcoal precipitation process formerly employed in Victoria 
m the chlorination process. 


The cyanide process has been successfully introduced in the states 
of California, Colorado, Idaho, Montana, Nevada, Utah, New 
Mexico, and Washington, and Black Hills district, South Dakota. 
Generally speaking, the adoption of the process in the States has 
been slow, the primary cause for this being doubtless due to the 
complex character of the ores. The progress of the process, how- 
ever, during the past year or two has been very marked. 

The process is one presenting many difficulties, especially with 
ores containing base sulphides, and up to the present time 
American metallurgists have been content to feel their way on safe 

The cyaniding of tailings is conducted on the same lines as else- 
where, while the slimes problem seems to have been successfully 
solved by the adoption of agitation followed by decantation. 

The treatment of the sulpho-telluride ores of Cripple Creek 
districts by cyanide after roasting has been attended with much 
silccess. Up to the present time, the electrical precipitation of 
gold from cyanide solutions has had little or no application on a 
working scale. For the treatment of high-grade slimes, there may 
possibly be scope for the filter-press process practised in Western 

The new Homestake plant, with a capacity of 1200 tons per 
day, and the Smuggler-Union of 600 tons per day, are among the 
largest in the States. Actual working details of cyanide treatment 
in the States are seldom available, and the following notes will be 
read with much interest. 


Colorado. — Camp Bird Mills, Ouray, — I am indebted to 
Godfrey Doveton for the following instructive particulars of the 
cyanide treatment of tailings at this mill. The tailings contain a 
small percentage of copper, which has led to the adoption of some 
interesting modifications of the usual practice. Mr, Doveton says 
the treatment of the tailings presents no new feature, but as much 
of the gold is partially enclosed in the coarser particles of sand, a 
rather lengthy treatment is required. 

The tanks are charged with Butters distributors, and the slimes 
overflow from three slime gates placed at the sides of the vats, 
and are conducted to the slime dam and settled. Slime gates are 
preferred to a circular launder, and as a somewhat better classifica- 
tion is effected, a better leaching is the result. 

A sizing test of a large number of vat samples showed the 
following results : — 


on 40 mesh, 

60 , 


80 , 


100 , 


120 , 








5 ] 

per cent 














The ore was crushed in stamper batteries through a 35-mesh 
wire-wove screen. 

On assay, the material found on the 40 and 60 meshes was 
found to run considerably higher than the finer material, and the 
bullion button was very much finer than that resulting from the 
assay of the finer product. 

The vats are sampled at the distributor nozzles as a check upon 
the vanner tailings at the stamp mill. When filled the charge is also 
sampled with a borer, some 40 to 50 bores being taken from a 500- 
ton vat. An acidity test for free and combined acid is made on 
the head sample, and the requisite quantity of lime is found, and 
added on top of the charge, and well mixed by shovelling over the 
top layer of material. 

Details of Treatment of 100-ton Charge. — The tank being partially 
drained and lime added, 20 tons of a weak solution are run on, 
containing about 0*05 per cent. KCy. This solution, which also 
contains a considerable percentage of copper cyanide, acts as a food 
for the cyanicides, and also, owing to the presence of much cupro- 
cyanide, dissolves much of the copper contained in the charge, 
thereby leaving the material in a better condition for the actual 


working solution. No gold is dissolved as a rule by the pre- 
liminary weak solution, and as the whole of the available cyanide 
is destroyed, it is allowed to drain off and run down the sluice 
launders, to be saved for sluicing purposes, etc. 

It should be mentioned that on rare occasions, when a portion 
of the gold contents is amenable to treatment, from 10 to 25 cents 
value of gold is removed by the first solution. 

The weak solution is immediately followed by a succession of 
10-ton charges of solution of 25 per cent, free KCy, until the out- 
going solution rises to about 0*20-0 225 per cent. KCy. Usually 
60 tons of solution, are run on, and by the time the outgoing 
solution is up to near standard strength, 65 to 70 per cent, of 
the gold is dissolved. The charge is now allowed to macerate 
from 8 to 16 hours, and is subsequently rapidly drained, sampled, 
and shovelled to the tank below for the second treatment. During 
the shovelling the charge sample has been assayed, and the gold 
yet capable of extraction ascertained. Usually almost all the 
soluble AuKCy 2 compound has been carried out by the charges of 
solution. Should the assay show that much of the previously 
insoluble gold still remains in the ore, standard solution is 
added in the lower vat. Should the gold contents be unusually 
high, 0*3, 0*35, or 0*4 per cent, solution is used for the saturation 
of the shovelled charge. ' If, however, a fairly good extraction has 
been obtained in the upper vat, 0*25 per cent, solution is used, 
and when we are crowded with tailings weak washes are applied 
immediately after the saturation. 

Mr. Doveton continues : — " Our usual practice is, however, to 
saturate with 0*3 per cent, solution, and allow to macerate 
from 4 to 6 hours, then drain rapidly, and apply a couple of 
washes of 0'3 per cent, solution of 10 tons each, drain partially 
dry, and give 25 tons of weak solution (O'08-O'l 25 per cent.). 
Before the last of the weak washes has disappeared below the 
surface, the charge is allowed to macerate two hours. This 
ensures all of the dissolved gold being carried out. 

" The charge is now drained dry on surface, and water washes 
added till cyanide values fall to 0*04 per cent. The liquor is now 
transferred from the weak gold tank to the waste gold tank, 
and the solution, low in cyanide, is passed through the waste zinc 
boxes, and thence out of the mill. The gold value of the solution 
at the time of transfer from weak to waste zinc boxes is about 
40 cents per ton, and much of this is precipitated in the 
waste zinc boxes. 

"At the time of sluicing, the charge is fairly dry, containing 
about 12 per cent, of moisture, and the outgoing solution 
usually contains about 0*02 per cent, of free KCy, but only a 


little double cyanide. As is usually the case, the first weak 
solution flowing from the charge is high in double cyanide, which 
we" are now about to recover by our recovery process. 

"It is noticeable that but little gold is carried out by the 
solution till the outgoing strength is almost up to standard 
strength of cyanide. 

"A series of solution assays were made on a number of vats 
during treatment. Samples of the outgoing solution were taken 
hourly, and the samples for each six hours assayed together. 
It was found that up to 0*05 per cent. KCy the solution 
carried scarcely any gold at all. Between 0*05 and 0*1 per cent, 
the value increased from 40 cents to 150 cents per ton ; from 0*1 
per cent, to 0*15 per cent KCy it was about 225 cents ; and at 
0*2 to 0*25 per cent. KCy the value remained about 5 dollars 
per ton of solution. The value seems to remain pretty constant 
till the solution falls to 0*2 per cent KCy, when a gradual 
decrease sets in, and at 0*05 per cent. KCy the value is usually 
80 cents per ton, and at 0*02 per cent KCy the value is seldom 
more than 25 cents. 

" There is a considerable quantity of copper in the solution, but 
nevertheless we have a most perfect precipitation. Our aim is 
always to precipitate the copper on the zinc with the bullion, thus 
keeping the copper contents of the solution at a constant figure. 

" Our strong sump assays for over a year have seldom gone as 
high as 40 cents, the average being from 15 to 20 cents. Our 
weak sump assays show but. an average of usually from 2£ to 4 
grains of gold per ton by the assay of very large samples, say 50 

"The precipitation is usually practically perfect the week 
following the clean-up. With our waste zinc boxes we are not 
quite so successful. The copper contents of the incoming solution 
are usually high, and the gold value only some 25 to 45 cents per 
ton. The available cyanide is very low, varying from a trace to 
•025 per cent., with the total cyanides 0*05 per cent. 

" By coating the zinc with mercury by immersing it in a weak 
solution of mercuric cyanide — made from mercuric chloride and 
KCy — we are enabled to make a very fair precipitation. 

" When the weak solution runs rather lower than usual in gold 
value, more of the copper is precipitated on the zinc, and a some- 
what less perfect precipitation of the gold takes place. The 
addition of concentrated cyanide solution at the head of the bin 
does not seem to help the precipitation of the bullion at all, but 
greatly accelerates the plating of the copper. The practice of 
strengthening the solutions entering the zinc boxes has been 
entirely discarded ; it has been found here most advantageous to 


promote the precipitation of the copper along with the bullion, 
thus preventing the solution becoming overcharged with copper. 
Occasionally, after placing fresh zinc in the last compartments of 
the boxes, it will be perfectly coated with copper in about eight 
hours, because the solution, when it reaches the last compartmetit 
or two, is very low in bullion value. However, in the course of 24 
hours, when perhaps a higher bullion value is entering the boxes, 
the gold will be plated over the copper, the precipitate being a 
very lustrous black. 

" Should the coppery zinc show no signs of precipitating bullion, 
precipitation is readily aided by the use of a few Winchester 
quarts of strong mercuric cyanide solution, which is allowed to 
drip into the compartment affected. When the zinc becomes 
gray in colour from the amalgamation of the deposited mercury 
with the coppery zinc, the addition of mercuric solution is dis- 
continued, and in a few hours the zinc will become black from 
bullion precipitate. 

" The evolution of hydrogen is much increased, should the solu- 
tions contain more copper than usual, and on several occasions, 
when the solution entering the boxes carried as much as 0*045 
per cent. Cu, and the usual bullion contents, 4 or 5 dollars per 
ton, the evolution of hydrogen was so great as to lift the zinc 
partially out of the boxes ; however, a good bullion precipitation 
was obtained, the solution leaving the boxes worth only 5 grains 
of gold per ton. An example of this was noticed recently, when 
a complete analysis of a working solution showed the presence 
of a small quantity of manganese, about 0*00778 per cent., 
estimated on the evaporation of 2 litres of solution by Volhardt's 
method. No sulphates or magnesia were found, but lime was 
present to the extent of 0*01387 per cent, of Ca. There was 
found a little lead, considerable alumina, no ferrocyanide, but a 
large amount of sulphocyanide, of the latter about 0*1078 per 
cent. There was a little iron present, but in what combination 
could not be determined. 

"No alkaline sulphide has yet been detected in the solutions 
here, and it would seem that when zinc is used as the precipitat- 
ing agent no alkaline sulphide can exist, the zinc being readily 
precipitated as sulphide from the ZnK 2 Cy 4 . 

" The oxygen is frequently estimated in the solution by a modi- 
fication of Threshis* iodimetric method. The estimation is very 
valuable, as it enables us to ascertain whether our solutions are 
sufficiently aerated or not. Numerous oxygen tests showed the 
presence of from 2 to 7*36 mgrms. of oxygen per litre of solution. 

" Experiments and numerous estimations were made to test the 
merit of employing jets of compressed air in the sumps and 


storages, should it be found that the solutions received insuffi- 
cient oxygen during their passage from the end of the zinc boxes 
to the leaching vats. However, we found that in the majority 
of the cases the solutions were capable of dissolving but a very 
little more oxygen on aeration, and the greater portion of that 
dissolved was diffused again on standing for any length of time. 

"The weak solution was found, on the average, to contain a 
little more dissolved oxygen than the strong, the increase being 
most noticeable when the weak solution leaves the zinc boxes. 
Here the weak solution containing 0'08 per cent. KCy and 0*11 
per cent, total KCy contained 0*64 mgrm. per litre of dissolved 
oxygen, while the strong solution contained only 4 mgrm." 

Cost of Cyanide Treatment at Gamp Bird Mills, 
5000 tons Tailings monthly. 

Cyanide, 1 lb. per ton, 
Zinc, 0*55 lb. „ 


)ents per ton. 

Lime, 1*8 lb. „ 

. 1-30 

Assays, . 
Labour, . 
Power — steam, 

3 60 


6 00 

Sand-pumps, maintenance, 
Electric light, . 
General repairs, 
Supervision, etc., 




Doveton gives the total charges as about 3s. per ton of ore 
treated, including 2*4 cents per ton for sulphuric acid treatment 
of gold slimes. 

Smuggler- Union Mine. — The cyanide plant at this mine was 
erected last year. It has a capacity of 600 tons per day. Zinc 
precipitation is used. 

Dorcas Pneumatic Cyanide Mill, Florence. — This plant has a 
capacity of 120 tons per day. The treatment here is of some 
interest, as compressed air is used to aerate and agitate the 
solutions and pulp in the leaching vats. The following descrip- 
tion is extracted from a paper by Dr. Wells.* 

The ore, averaging 20 dollars in value, passes through crushers 
and coarse rolls to the sampling room, thence through a revolv- 
ing dryer and through two sets of finishing rolls which crush it to 

* The Engineering and Mining Journal, Jan. 4, 1902. 


2 4 -mesh size. The finely crushed ore is then roasted in a Holt- 
hoff-Wethey furnace, the roasted product passing to the leaching 

There are six of these tanks, 30 ft. in diameter, 4*5 ft. deep, and 
fitted with air pipes in the bottom for the introduction of air 
during the leaching. The air is supplied at a pressure not 
exceeding 5 lbs., about 1 cubic foot of air per minute to each ton 
of ore being sufficient for agitation and oxygenation. 

The treatment generally lasts five days, and leaves a value of 
about 1 dollar per ton in the ore. The residue is then sluiced 
out and concentrated on Wilfley tables. The tailings after this 
treatment average about 40 cents per ton. 

The tanks are filled by a conveyor, and as soon as the bottom 
of the tank is covered with pulp a solution containing 10 lbs. 
KCy per ton is run in gradually, the pulp at the same time 
continuing to flow in until the tank is full. The air is then 
turned on gradually, and is kept on until the pulp shows an extrac- 
tion of at least 90 per cent. Whenever the air comes up un- 
evenly through the charge the ore has to be stirred by men with 
iron rods. This generally requires about 30 hours. The air is 
then shut off and the pulp allowed to settle for one hour. 
Percolation is then begun and the strong solution run off as 
quickly as possible, followed by a weaker one of 5 lbs. KCy per 
ton. This operation is continued until the tests of the solution 
show only traces of gold. Water is then added to displace 
the KCy solution. The loss in cyanide is stated to be less than 
1 lb. per ton. The dust resulting from the dry-crushing is 
collected and treated with cyanide, without previous roasting. 
From dust assaying 51 '20 dollars per ton the tailings only contained 
80 cents per ton. It is added to the roasted ore in the leaching 
tanks, 3 tons to each tank, spread evenly on top of the charge. 

Cripple Greek Telluride Practice. — The country rock of this 
district is andesite breccia, phonolite, and decomposed granite. 
On the surface the ores are oxidized and carry iron peroxide, 
manganese oxide, and oxide of tellurium. Below water-level, the 
gold occurs in the. minerals calaverite and sylvanite, and is always 
associated more or less with iron pyrites. 

The gold in the surface ores is free, but not amenable to 
amalgamation, being coated with metallic oxides. It is, however, 
easily extracted by cyanide. 

The unoxidized telluride ores have to be subjected to a pre- 
liminary dead roast before cyanide treatment. 

The reduction of the ore is generally effected with Krom rolls, 
instead of the mills used at Kalgoorlie, the ore being crushed to 
pass through a 40-mesh screen. The roasted ore is leached by 


percolation in vats with two solutions of cyanide! The stronger, 
containing from 0*5 to 0*75 per cent, of potassium cyanide, is 
allowed to percolate for 50 hours, after which the weak solution 
is added. The time of treatment varies from 70 to 100 hours. 

Precipitation is effected with zinc* 

The treatment of the sulpho-telluride ores of Western Australia 
is obviously a more difficult problem than that of Cripple Creek 
ores. At Kalgoorlie, the ore is talcose and is highly calcareous, 
and forms so large a proportion of slime that ordinary percolation 
is impossible, and deoantation too slow and costly. The success- 
ful solution of the difficulty was found in the use of filter-presses, 
in which the dissolution of the gold is effected under pressure, and 
by means of which the slimes are easily separated from the gold- 
containing solutions and washes. 

As at Cripple Creek, the oxidized ores are treated in the raw 
state, while the undecomposed ores are subjected to a dead roast 
before cyanide treatment. With the adoption of filter-presses, it 
was soon recognized that the greater the proportion of slimes the 
better, and to attain that object, pulverizing mills, instead of rolls, 
have come into general use. 

California. — California King Mines. — The plant at this mine 
has a capacity of 1000 tons per day. The ore is crushed by rolls 
to 20-mesh size and then treated in ten cyanide vats, each 40 ft. 
in diameter and 5 ft. deep, provided with bottom-discharge doors. 
Zinc precipitation is used. 

There are many cyanide plants scattered throughout the 
State, including the large 140 -stamp cyanide mill of the 
Golden Cross Mines at Hedges, but no particulars of these 
are available. 

Montana. — Dr. Wells says that cyaniding has made great 
progress in this State. Many new plants have been erected, some 
of them of large size. 

The process as carried on in this State is much the same as 
elsewhere. In many of the mills, the method is wet-crushing, 
amalgamation on plates, followed by cyaniding of tailings. In 
other cases the ore is crushed in rolls and cyanided direct. With 
slimy ores, the practice is to agitate and decant. 

Nevada. — The sun-drying slime treatment adopted at the 
Dexter plant at Tuscarora has already been referred to in the 
chapter dealing with slime treatment. Of the other plants in 
operation in this State some are treating tailings, while others are 
dealing directly with dry-crushed ore. 

New Mexico. — The Cochiti Company of Bland has a successful 

* Prof. Furraan. Mines and Minerals, January 1897. 



method of treatment which possesses some novel features.* Here 
98 per cent, of the dust will pass a 100-mesh screen. The process 
used is agitating and settling, the agitation being effected by 
steam and compressed air, the latter being under 60 -lb. 
pressure. Steam is used to heat the charge, as it was found that 
the air on expanding cooled it too much. The mixture of air and 
steam is admitted through 1 in. horizontal iron pipes in the 
bottom of the tank. The pipes have 0*0625 in. holes in the under 
side for the escape of the air and steam. In 24 hours 94 per cent, 
of the total values are in solution, using a solution of cyanide of 
3 lbs. per ton, but only 80 per cent, is recovered, due to the 
solution going to waste in the slimes. To neutralize the acidity 
of the ore and to aid the settling of the slimes, lime is suspended 
in a wire basket in the upper part of the tank. The cost of 
treatment is given as 1 dollar per ton. 

Utah. — Mercur is one of the greatest centres of the cyanide pro- 
cess in the United States, the different plants having a capacity 
of over 2000 tons per day. The Mercur ore has been described as 
follows t : — " Silica in a form similar to silicious sinter, or gey serite, 
characterizes the ore. Cinnabar is most abundant in this rock, 
and forms beautiful incrustations in the cellular varieties. 
Wherever found in the district, cinnabar is considered a sure sign 
of gold. Orpiment and realgar occur in large quantities in some 
of the ore. There is about as much iron as is usually found in 
impure limestone and clay. Barite and gypsum occur more or 
less crystallized : also masses of limestone are found mineralized 
in rings, the outside assaying from 6 dollars to 8 dollars, and the 
centre a trace in gold. No free gold is visible in the ore, even 
with a microscope. One remarkable feature is the absence of 
silver. The average of the ore milled is kept close to 1 2 dollars 
per ton." 

At the Mercur mill the ore is delivered by the railroad to an 
ore-bin 40 ft. long, 20 ft. wide, and 20 ft. deep. It is crushed in 
a Dodge crusher, from which it passes to a set of Walls corrugated 
rolls, and is finally trammed to the cyanide vats, after being 
crushed to 1 in. mesh or less. The vats are 12 ft. 8 in. in dia- 
meter, and hold 15 tons. They are made of tank iron, with red- 
wood bottoms. The filter cloth on the false bottom is burlap, and 
lasts from four to six weeks. From the tanks the solution is con- 
veyed to a collecting tank, from which it is pumped by Blake 
single-acting pumps to the precipitating room. The zinc boxes 
are from 24 to 36 in. wide, 10 to 12 in. deep, and about 20 ft. 

* Hunter, Engineering and Mining Journal, Jan. 19, 1901. 
t Mining and Scientific Press, May 23, 1896. 


Fine crushing is found to be unnecessary, as the ore is very 
porous, and much of it disintegrates into mud when solutions are 

It is interesting to note one change made in Mercur practice. 
Formerly the strong solution was run through the ore continu- 
ously, the surface being kept always covered. Now a series of 
washes is run through, the solution each time being down below 
the surface. The extraction has been increased thereby, and 
much time saved on each vat. 

The solution used is from O'l per cent, to 0*3 per cent, in strength. 
It was at one time the practice to estimate the strength of the 
solution by its action in the zinc boxes, and by its alkaline feel. 
At the present time more acute methods are practised. Still an 
instance has come within our notice of an operator determining 
the strength of his solutions wholly by their odour. 

At the Mercur mill the practice was formerly to standardize 
solutions by adding cyanide to the lower end of the zinc boxes, 
" the judgment of the operator determining the time and amount." 

The zinc slimes were dried in an old retort belonging to the 
amalgamating mill. The door is closed but not luted, and at 
about 160° C. the product ignites, producing fumes of a complex 
nature, causing salivation and a headache. The slimes are finally 
taken from the retort, and the burning completed on a sheet-iron 
table. This product is then shipped to a smelter for refining. 

At present the residues from a 12 dollar ore assay about 1*75 
dollars, giving an extraction of 85 per cent. The average ore 
value is about 6 dollars per ton. The cost of treatment is itemized 
as follows : Mining, 35 c. per ton ; railroad hauling and milling, 
80 c. ; cyaniding the ore, 1 dollar 35 c. — total cost per ton, 2 
dollars 50 cents. The consumption of cyanide is at present about 
\ lb. per ton of ore. 

TJie Golden Gate Cyanide Works. — These are said to be the 
largest and best equipped in the United States.* They were 
constructed in 1898, and are built on a hillside with eight levels. 
In order to get the ore to the top of the works it has to be 
hoisted on an incline 800 ft. long. The mill is 294 ft. wide, and 
420 ft. in length up and down the slope. The difference in 
elevation from top to bottom is 145 ft. The retaining walls, 
which are 2 ft. wide at the top, and have a batter of 1 ft. in 12, 
required over 50,000 cb. yards of rubble masonry. The various 
floors were constructed by blasting out the side hill. The broken 
stone thus obtained was used for the retaining walls and filling 
behind them. The mill is driven by power transmitted electri- 

* Excerpt from Bosqui's Cyanide Process, p. 174. 


cally a distance of 35 miles at a tension of 40,000 volts. The 
loss of energy in transmission is said to be only 5 per cent. 

At the works the 40,000 volt 3-phase current is transformed to 
one of 220 volts of 2-phase. The current is delivered at a con- 
tract price of 60 dollars per H.P. The first section of the mill 
contains the coarse crushers, and in the second are the dryers. In 
the third section is the fine crushing machinery, which consists of 
four sets of 26 -in. rolls, and three sets of 36 -in. Berthelet 
apparatus are used for sizing. There are six elevators, with a lift 
of 60 ft. The fourth, fifth, and sixth sections contain the roast- 
ing furnaces, which are of Brown's straight-line design, four in 
number. Those intended for arsenical ores are estimated to have 
a daily capacity of 75 tons, while those for talcose ores are rated 
at 150 tons. The ore is stirred by the rabbles once each minute. 
One man attends to two furnaces. The gases are carried from the 
furnaces through 6 by 8 ft. flues into the main dust chamber, 
which connects with a steel stack, 8 ft. in diameter and 85 ft. high, 
located on the hill above the buildings. The top of this stack is 
275 ft. above the lowest level of the building. The leaching 
department, which constitutes section 7, is 60 by 294 ft. It has 
two floors, the main floor supporting ten tanks 25 by 50 ft. and 
5 ft. deep (presumably rectangular), and three solution tanks 20 ft. 
in diameter and 12 ft. deep. The tanks are supported by masonry 
piers. They are charged by hand from cars run on bridges over 
the tanks. The eighth section of the mill, which is 50 by 70 ft. 
and two stories in height, is the precipitation department. It 
contains three precipitation tanks, each 14 ft. in diameter and 
8 ft. deep. The tailings from the leaching tanks are discharged 
into cars which are run to the waste dump. The building is con- 
structed of steel. 

South Dakota. — The following details of cyanide plants and 
practice are extracted from a paper contributed by Mr. Chas. H. 
Fulton, M.E., to the Engineering and Mining Journal, January 
4, 1902. 

The Homestake Company's 1200-ftm Tailing Plant at Lead, — 
This plant takes the tailings pulp from the Golden Star, Eighty, 
and Highland stamp mills. The pulp is conveyed to the cyanide 
plant in a 10 in. cast-iron pipe. On its way it passes through the 
classifier house, where some of the slimes are separated by means 
of eight large sheet-iron cone-shaped classifiers. These cones have 
no upward or rising current, but as the ore pulp is charged by pipe 
at the centre, the sands settle to the bottom and are discharged, 
while the slimes overflow at the periphery. The sands are then 
carried to the cyanide plant, where they are fed to seven addi- 
tional cones and treated as above described. The slimes overflow- 


ing at the periphery are again discarded, while the sands are dis- 
charged to four-compartment jigs of the Harz pattern, which act 
as classifiers. One cone feeds four jigs. The jigging is done on 
beds of pyrites concentrates, relatively coarse screens being em- 
ployed, except in the last compartment. The stroke varies in the 
different compartments from £ in. to about £ in. The last com- 
partment is widened out in order to compensate for the increased 
amount of discharge water which accompanies the slimes during 
their carriage from the tanks. The continuous hutch discharge is 
taken by open launder to the revolving pipe distributors, which 
fill the leaching tanks, the peripheral overflow of the tanks still 
carrying off slimes. Lime in the form of an emulsion is added to 
the pulp launder which takes the jig discharge. 

There are fourteen leaching tanks, 45 ft. in diameter, 9 ft. deep, 
built of California red- wood. The tanks are placed seven in a row. 
There is one revolving distributor of the Butters type for each 
row, being moved from tank to tank by overhead track, as required. 
To get an even overflow at the periphery of the tank, the tops of 
the staves are grooved to a depth of £ in., and a soft pine tongue 
is inserted, projecting from \ to J in. above the top of the staves. 
This is then easily planed, and can be kept perfectly level. It is 
also readily renewed. 

It is the object to charge just as much slime with the sands as 
possible, and still leach in a reasonable time. For this purpose 
the adjustments are made on the jigs and distributors. The 
value of the pulp as it leaves the stamp mills is from 1.25 dollars to 
1.50 dollars per ton. The amount crushed in the mills approxi- 
mates 2000 tons per day, while the cyanide mill treats about 1200 
to 1300 tons per day. The difference, or 700 tons, is discarded as 
slimes which, while carrying considerable value, cannot be treated 
in the condition they are in, and must be discarded. The propor- 
tion of slimes separated out is about 35 per cent. The sands are 
leached directly in the filling vats. 

The Cyanide Plant of tlie Wasp No. 2 Mining Company. — This 
plant on Yellow Creek, near Kirk, is a dry, coarse crushing plant, 
capacity 90 tons per day. The ore is quartzite, the gold being 
on the cleavage planes, and in the seams and cracks, so that coarse 
crushing only is necessary to make it available for solution. The 
value varies from 3 dollars to 12 dollars per ton, the main bulk 
averaging about 4 dollars. It is mined very cheaply (85 to 90 
cents) by open-cut mining. The mill is built in terraces on a very 
steep hillside, and the ore is moved through almost entirely by 
gravity. It is hauled from the mine 750 ft. distant in trains of 
four cars, and charged into the main bin, from which it passes to 
a No. 2 D. Gates crusher over a 4 by 8 ft. grizzley having 1 J in. 


spaces between bars. The undersize and the crushed ore from the 
Gates go to the storage bin, from which it is fed by Tulloch feeders 
into one set of coarse Gates rolls 14 by 24 in., 80 revolutions 
per minute. The rolls discharge to a stationary inclined 2-mesh 
screen 7 ft. long, 1 ft. wide, the oversize from which passes 
through the fiuishing rolls 14 by 24 in., operated at 100 revolu- 
tions per minute, and meets the undersize, the combined product 
being elevated by the one-bucket elevator in the mill to the shak- 
ing finishing screens, situated on a level with the coarse rolls. 
This screen is an inclined shaking 2 J-mesh screen, 1 6 ft. long and 
2 ft wide, the lower half being stationary. Nearly all of the 
finished product passes a 6-mesh screen, the oversize from which 
is returned by gravity to the finishing rolls. 

The ore is charged from the finished product bin into the four 
leaching tanks of Oregon fir; each tank is 16 ft. diameter, 7 ft. 
deep, and of 55 tons capacity. For this purpose a 16 in. belt con- 
veyor is used at a speed of 600 ft. per minute. A large 100-ton 
tank has recently been added, which gives the mill the capacity 
mentioned above. Tanks are charged in from two to two and a 
half hours, the cyanide solution being added after the tank is 
about one-half filled. The strength of the strong solution is 6 
lbs. cyanide per ton of solution. About 15 lbs. of strong solution 
are run on, and usually allowed to stand three to four hours, 
then it is drained, a weak solution being added at top to replace 
it. After this complete replacement it is allowed to stand a 
short time and then circulated for the balance of the leaching. 
The weak solution contains from 2 J to 3 lbs. cyanide per ton. 
About 50 tons of weak solution are used, followed by from 6 to 9 
tons of wash water. To neutralize the acidity of the ore and give 
a protective alkalinity, 6 lbs. of lime per ton are added before 
the ore is dumped into the first bins. The cost of cyaniding is 
from 85 to 90 cents per ton. The precipitation is accomplished 
by zinc thread in specially constructed barrels, and very good 
results are obtained, the tailings solution having a value of from 
4 to 10 cents per ton. The extraction made is from 80 to 90 
per cent, of the values, and the consumption of the cyanide 
amounts to about \ lb. per ton of ore. The precipitates are 
treated by the usual sulphuric acid method. Zinc dust precipita- 
tion in the tanks, pumping the resultant mixture through filter- 
presses, was tried at the mill, but discarded on account of the 
difficulty in the clean-up. 



Athabasca Mine, Nelson. — In a paper read before the Mexi- 
can meeting of the American Institute of Mining Engineers in 
1901, Mr. E. Nelson Fell gives some interesting information on the 
treatment of tailings at this mine. The ore, he says, consists of a 
quartz gangue, containing a little lime and variable quantities of 
the sulphides of iron, lead, and zinc. The following figures, giving 
the analyses of the ore before milling, and of the tailings after 
milling, which constituted the material to be cyanided, are based 
on the daily samples taken during February 1901. 

Analyses of Ore and Tailings. 





Per cent. 

Per cent. 



7 04 












Oz. per ton. 

Oz. per ton 










A1 2 3 , 



The analysis shows this to be an ore well adapted for cyanide 

The plant was located to receive the tailings direct from the 
mill, in two distributing- tanks, 14 ft. in diameter and 10 ft. in 
height. The tanks are fitted with annular launders around the 
rim, and are filled with water before the admission of the tailings ; 
while the overflow is carried off in annular launders. 

A slimes-arrester is fixed in each vat, consisting of a sheet of 
iron, 10 in. wide, fitted inside each vat, about 1 in. from the staves, 
and extending all the way round. This ring of sheet iron is thus 
2 in. less in diameter than the diameter of the vat, and is held in 
position by eight iron brackets so arranged that the ring can be 
raised entirely above the level of the tank, or lowered until the 
upper edge is about level with the top of the staves, according to 
the nature of the ore being treated. The obvious intention of 
using this contrivance is to retain as much of the slimes with the 
sands as possible. 


The slimes-arrester described by Mr. Fell is exactly the same as 
that attached to the J. C. Fraser and Price's continuous grinding 
and amalgamating pans, which were extensively used in Australia 
and New Zealand for the treatment of mill tailings before the 
introduction of the cyanide process. 

The overflow takes place in the space between the ring and the 
staves, and having such a great length of outflow, is very even and 

The plant, which has a capacity of 50 tons per day, consists of 
five leaching- vats, each 18 ft. in diameter and 4 ft. deep, fitted 
with centre discharge doors, and two collecting- vats, each pro- 
vided with the ordinary revolving reaction distributors. The latter 
are 14 ft. in diameter and 10 ft. deep, and so situated above the 
leaching- vats, that each one can be discharged by shovelling 
through side-doors into any one of three adjacent leaching-vats. 

Besides these there is a strong gold tank, a weak gold tank, a 
waste-water vacuum tank, and two series of zinc extractors, each 
consisting of twelve iron boxes, which are square and have each a 
capacity of one cubic foot of zinc-shavings. The boxes are movable, 
and each is independent of the other. The two gold tanks are 10 
ft. in diameter and 6 ft. deep, and the two sumps 12 ft. in dia- 
meter and the same depth. 

Mr, Fell supplies the following working details : — 

" This description shows the actual details of treatment, and the 
results obtained from the treatment of Lot 59, which gives a fair 
example of the treatment. 

" June 26th, at noon, turned on strong solution (0*24 per cent.) 
until same stood 6 in. deep on the sands ; allowed to stand 4 
hours. At 4 p.m. opened outlet-cock and allowed solution to run 
into gold-tanks ; at 4.30 p.m., as the last of the solution was pass- 
ing off, took sample, which assayed "nothing" in cyanide and 
9*30 dollars in gold. Closed outlet-cock, admitted fresh charge of 
strong solution, and allowed to stand 8 hours. At 12.30 (mid- 
night), June 27th, opened outlet-cock, and at 1 a.m. took sample, 
as before, which assayed 0*06 cyanide and 28*94 dollars in gold. 
At 1 a.m. turned on weak solution (0*08 cyanide), allowing same 
to drain through without interruption till 4 a.m. Sample at 4 
a.m. assayed 0*10 cyanide and 8*08 dollars in gold. Shut off 
outlet-cock and allowed solution to stand until 1.30 p.m. Opened 
outlet-cock and admitted fresh solution (0*06 cyanide) and allowed 
same to run through, admitting fresh solution as required till 
12.30 (midnight), June 28th. Sample taken at 10 p.m. assayed 
0'07 cyanide and 0*62 dollar in gold. Closed outlet-cock and 
allowed to stand until 5 a.m. Opened outlet-cock at 5 a.m. and 
allowed fresh solution to run through until 11 a.m. Sample 


taken at 6 a.m. ran 0*06 cyanide and 0*42 dollar in gold; and 
sample taken at 11 a.m. ran 0*6 cyanide and 0*21 dollar in gold. 
At 11 a.m. turned in water-wash till 2 p.m. Sample taken at 
12.30 p.m. ran 006 cyanide and 0*21 dollar in gold; and sample 
taken at 2 p.m. ran 0*06 cyanide and 0*21 dollar in gold. At 
this point the process was declared finished ; the wash was drained 
to waste ; and the tailings were discharged. 

"Assay of the tailings before treatment gave 13*02 dollars, and 
after treatment 2*07 dollars in gold. Percentage of recovery, 
84-1. Time occupied, 2 days 2 '5 hours. " 

If the actual recovery is equal to the extraction obtained by the 
assay difference, the results obtained must be considered very 
satisfactory, and with prolonged leaching could doubtless be 


Laurence Pitblado states* that in the Kolar field there are at 
present six cyanide works in operation. The ores of the field are 
very simple, consisting mainly of pure quartz, with only a small 
percentage of iron pyrites. The material treated is tailings from 
the stamp mills. Those first worked in the Mysore plant averaged 
4*5 dwt. gold, and yielded 65 per cent., with an average consump- 
tion of 1 lb. cyanide per ton. In 1897 a month's test with 40-mesh 
screens in the batteries gave the following result : — 90*65 per 
cent, amalgamated in batteries and on plates ; 74 per cent, of the 
value in the tailings recovered by cyanide lixiviation, making a 
total extraction of 97 per cent, of the ore as delivered to the mill. 
The cost of cyaniding at the present small plant of the Mysore 
Company is £2, 10s. 5d. per ton, exclusive of royalty and de- 
preciation, but in the new 4000-ton plant which is being erected 
alongside the heap of accumulated tailings that is to be worked, it 
is believed that the cost will not greatly exceed 2s. At the 
Champion Reefs mill, where 20-mesh screens are used in the 
batteries, the extraction from the tailings is about 56 per cent, 
with a consumption of 1 lb. of cyanide per ton. 

In refining the precipitate in this district it is first passed 
through a 30-mesh screen, drained, dried, roasted with or without 
a small percentage of nitre, and fused directly in plumbago 
crucibles. At the Mysore works the precipitate is retorted before 
roasting, yielding about 100 lbs. of mercury per month. The 
presence of mercury in the zinc boxes generally leads to the pro- 
duction of much-floured and brittle zinc. In fluxing and smelting 

* Journal of the Society of Chemical Industry, Feb. 28, 1898. 


directly, the retorted slimes the bullion assayed 56*4 per cent, 
gold, 3 per cent, silver, 2*4 per cent, lead, 19*6 per cent, copper, 
18 per cent, zinc, and 0*1 per cent, nickel. The slag contained a 
good many shots of metal. In roasting with nitre a slag freer than 
the above was produced, and a bullion assayed 49 '5 per cent, gold, 
4*6 per cent, silver, 4*29 lead, 38*21 per cent, copper, 2*10 zinc, 
and 0*22 per cent, nickel. These results being unsatisfactory, the 
following practice was adopted at the Mysore works. The retorted 
and dried slimes are mixed with 10 per cent, of nitre and roasted 
at a bright red heat. When cold they are boiled with dilute sul- 
phuric acid (1 : 2), which dissolves the copper. The dried and 
washed product is fluxed with about 35 per cent, borax, 15 per 
cent, soda, and 10 per cent, sand, giving a slag free from shots of 
metal and a bullion assaying 81 *3 per cent, gold, 6 '9 per cent, 
silver, 2*71 per cent, lead, 6*78 per cent, copper, 0*4 per cent, 
zinc, 0*12 per cent, nickel. 



The distinguishing features of this process are the use of extremely 
dilute solutions of cyanide and the electrical precipitation of the gold. 

Since the introduction of the cyanide process, the precipitation 
of the gold by metallic zinc has always been regarded as a weak 
point ; and metallurgists have devoted much time in the endeavour 
to discover an efficient substitute for it. 

Electrical precipitation naturally engaged the attention of many 
investigators. In 1893, the author, assisted by Mr. F. B. Allen, 
M.A., B.Sc., conducted a number of experiments with electrical 
precipitation to determine the method of precipitation to be 
adopted at the School of Mines cyanide plant. Many different 
modifications were tried. With some, the precipitation from 
solutions of ordinary working strength was very satisfactory ; but, 
with all, the precipitation of the gold from dilute solutions, such 
as those corresponding to weak cyanide washes, was always very 
imperfect and accompanied by decomposition of the water. 

In the Siemens-Halske process this difficulty is overcome by 
causing a slow artificial circulation of the cyanide solutions in the 

The plant and operations connected with the leaching of the 
gold are the same as those described in the preceding chapters, 
the only difference being in the extractor-house. 

The electrical precipitation of gold has been introduced with 
marked success at a number of cyanide plants at the Witwaters- 
rand Goldfields, and its use is extending. Up to the present it 
has not been introduced to New Zealand or Australia, and so far 
very little has been written about it. For the following details of 
the process I am indebted to the papers of Mr. Charles Butters 
and Mr. A. Von Gernefc, read before the Chemical and Metallur- 
gical Society of South Africa, and published in the South African 
Mining Journal. 


Discovery of the Process. — Mr. Von Gernet said the 
electrical precipitation of gold extracted from ores by cyanide 
has been in use in Europe and Asia as far back as 1888. In 
1887, Dr. Siemens found that the gold anodes used in electro- 
plating at his works in Berlin lost weight when standing idle in 
the cyanide solution, without any electric current passing through 
the bath. This, in connection with the well-known fact that gold 
was soluble in aqueous solutions of cyanide, first induced him to 
try the use of that solvent for the extraction of gold from ores. 

In the same year he built a small plant to make experiments 
on concentrates produced in Siebenburgen. The gold was pre- 
cipitated both by electrolysis and zinc filings. It was found, 
however, that the zinc method gave good results only from com- 
paratively strong solutions, while the electrical precipitation was 
effected with both dilute and strong solutions, and its efficiency 
was not affected by the presence of caustic soda. 

Dr. Siemens therefore decided to use electrolysis only, and early 
in 1888 he commenced operations on a large scale. Engineers 
were sent to different countries, two going to Hungary, one to 
America, and one (Mr. Von Gernet) to Siberia. 

The operations were generally successful, and in May 1894, a 
plant, capable of treating 3000 tons of tailings per month, was 
erected at the Worcester mine, near Johannesburg. During 1895 
the process was adopted by some eight or ten large mining com- 
panies, including the Metropolitan, May Consolidated, Croesus 
G.M. Co., No. 4 Central Works, and Robinson Slime Works ; and 
already it is a formidable rival of the M c Arthur-Forrest zinc- 
precipitation process. 

Action of the Electric Current on the Cyanide Solu^ 

tion of Gold. — The electric current decomposes the auro-potassic 
solution, depositing the gold on the negative pole and liberating 
the metalloid at the positive pole. In a fixed time a given 
electric current will deposit a certain quantity of metal, which 
quantity will vary for different metals in direct proportion to 
their electro-chemical equivalents. This law holds good only for 
solutions strong in metal ; but with very dilute solutions, as in 
use in the cyanide process, the current does not find sufficient of 
the metallic compound present at the electrodes, and consequently 
decomposition of water also takes place ; for this reason, to make 
the precipitation as efficient as possible, constant diffusion of the 
solution is required. 

The artificial circulation of the solution is most economically 
and conveniently obtained by allowing a slow but steady flow 
through the precipitation boxes. It is of the highest importance 
to give a very large surface to the electrodes, since a more efficient 


precipitation is obtained by doubling the number of plates than 
by increasing the current tenfold. 

The Cathode or Negative Electrode. — To obtain a satis- 
factory cathode, a metal must be used which will fulfil the follow- 
ing conditions : — 

1. The precipitated gold must adhere to it. 

2. It must be capable of being rolled out into very thin 

sheets to save unnecessary expense. 

3. It must be easy to recover the gold from it. 

4. It must not be more electro-positive than the anode, in 

order to prevent return currents being generated when 
the depositing current is stopped. 

The most suitable metal was found to be lead, which, in the 
form of foil, meets all the requirements, and is therefore used in 
the Siemens-Halske process. 

The Anode or Positive Electrode.— The requirements 

of the anode are no less important. By the action of the current 
a metalloid is liberated at the positive electrode, and the latter, 
when a metal, begins to oxidize. Carbon could be used, but it 
will not withstand the action of the current, soon crumbling into 
a powder which decomposes potassium cyanide. Besides, when 
this finely-divided carbon is in suspension, it cannot be removed 
from the solution by filtration. 

When zinc is used as an anode, it forms a white precipitate of 
ferro-cyanide of zinc by the reaction of zinc oxide upon ferro- 
cyanide formed during leaching. In the same way, iron anodes 
form Prussian blue by the reaction of oxide of iron and ferro- 
cyanide. In consequence of this reaction, the amount of ferro- 
cyanide in the cyanide solution does not increase. 

The cyanide can be recovered from the Prussian blue, formed 
at the iron anodes, by dissolving it in caustic soda, then evaporat- 
ing the solution, and finally smelting with potassium carbonate. 
Mr. Von Gernet states that this process has been tried on a small 
scale, about 50 lbs. at a time, with the result that a nice clean 
potassium cyanide was obtained. In the treatment of clean 
tailings, this regeneration of cyanide is not of great importance ; 
but with concentrates, or tailings, which decomposes the cyanide 
solutions with formation of ferro-cvanide, it will effect a consider- 
able saving. 

Electric Current Required for Precipitation. — Only a 

very weak current is required to precipitate the gold from cyanide 
solutions, a density of 0*05 ampere per square foot being sufficient. 
With cathodes 1J in. apart, 7 volt is sufficient to produce this 
strength of current. 


The advantages gained by using such a weak current are : — 

1. The gold is deposited hard on the lead- foil. 

2. The iron anodes are preserved for a long time, as their 

waste is in proportion to their current strength. In 
a plant treating 3000 tons per month, 1080 lbs. of iron 
are destroyed in that period. 

3. Little power is required. 746 Watts equal 1-horse 

power. A 3000- ton plant requires 2400 Watts, equal, 
theoretically, to 3|-horse power, and actually requiring 
about 5-horse power. 

The Advantages of Electric Precipitation.— The prin- 
cipal advantages claimed for this process are as follows : — 

1. That the precipitation operates independently of the 

amount of cyanide or caustic soda present in the solu- 
tion. Therefore, in the treatment of tailings, very dilute 
solutions can be used, the only limit being a sufficient 
quantity of cyanide to dissolve the gold satisfactorily. 
A solution containing 0*03 per cent, of cyanide will dissolve 
gold as effectively as a solution containing 0*3 per cent., provided 
a longer time is allowed for treatment. In the first case, the 
decomposition of cyanide is much less than in the second, result- 
ing in a corresponding economy. 

2. However acid the solution may be when entering the ex- 

tractor, the precipitation takes place equally as well 
as it does when the solution is neutral or alkaline. 

3. No complications arise from the formation of lime, alumina, 

or hydrate of iron, which sometimes cause trouble in the 
zinc process of precipitation. 

4. With ores or tailings containing copper, the extraction of 

the gold will be the same, but the decomposition of 
cyanide less than when using stronger solutions. 

5. The successful treatment of slimes. 

The Actual Working of the Process. — The first practi- 
cal demonstration of this process on a commercial scale took 
place at the cyanide works of the Worcester Gold Mining Com- 
pany, near Johannesburg, under the supervision of Mr. A. Von 

The plant consists of five leaching vats placed on a row of 
stone piers, with a single tunnel beneath. Each vat is 20 ft. 
in diameter, with 10 ft. staves, and has a capacity of 100 tons 
of tailings. 

Between the vats and the electric extractors there are placed 
two tanks, 16 ft. in diameter, with 6 ft. staves, forming two 
intermediate reservoirs, which enable the flow through the pre- 


cipitation boxes or extractors to be kept constant and steady, a 
matter of great importance. 

A better method to secure an even flow is to pump all the 
solution into a small raised tank, provided with an overflow into 
the intermediate tank and a delivery pipe to the precipitation 
boxes. The small tank is always kept full to overflowing, so that 
it delivers under a constant hydraulic head. 

Beyond the precipitation boxes there are two sumps, 20 ft. in 
diameter and 6 ft. deep, from which the cyanide solutions are 
returned to the leaching vats. 

Two collecting vats, 20 ft. in diameter and 8 ft. deep, receive 
the tailings from the 25-stamp battery. 

The Electric Precipitation Boxes. — There are four pre- 
cipitation boxes, constructed of wood, each 18 ft. long., 7 ft. wide, 
and 4 ft. deep. Each box contains 89 iron-plate anodes, 7 ft. by 
3 ft. by £ in., cased in canvas to retain the small quantity of 
Prussian blue produced ; and 88 cathodes of lead-foil stretched on 
iron wires fixed on a wooden frame. Each frame contains three 
strips, 3 ft. by 2 ft., so that, counting the double surface of each 
lead-sheet, there are altogether about 3000 square feet of cathode 
surface, the current density being 0*05 ampere per square foot. 
Copper wires are fixed along the top of the sides of the boxes, and 
convey the current from the dynamo to the electrodes. 

The boxes are made of 3 in. material throughout, with stiffen- 
ing pieces across the sides and bottom. The divisions are of wood, 
or are formed by raising some of the iron plates about an inch 
above the level of the solution, while others rest right down on 
the bottom, the joints being made water-tight by means of wooden 
fillets caulked with hemp packing. By this means a series of 
compartments is obtained, similar to those in a zinc precipitation 
box, the difference being that the solution passes alternately up 
and down through successive compartments. The rate of flow is 
about one foot per minute. 

The Clean-up. — The boxes are kept locked, being opened 
once a month for the "clean-up," which is conducted as follows: — 
The frames are taken out singly, and the lead-foil is removed and 
replaced by fresh lead-foil, the whole operation taking but a few 
minutes for each frame. The lead, which contains from 2 to 1 2 
per cent, of gold, is then smelted into bars and cupelled. 

The gold is deposited on the lead sheets as a thin bright yellow 
film, which adheres firmly to the lead. The consumption of lead 
at the Worcester Works is 750 lbs. per month, equal to 1 £d. per 
ton of tailings. 

The working expenses for treating 3000 tons per month were as 
follows : — 




Filling and discharging leaching vats, . 

, 10-00 per ton. 

Cyanide, ^-lb., 

, 600 „ 

Lime, .... 

. 1-20 


Caustic soda, . 



Lead, ...... 

. 1-10 , 


Iron, ...... 



White labour, .... 



Native labour and food, 

. 1-90 , 


Coal, ..... 

. 4-60 , 


Stores and general charges, 

, 3-30 , 


Total, . 

, 36 00 per ton 

of 2000 lbs. 

The cost of treatment per ton of 2*240 lbs. would be 3s. 4*32d. 

The tailings assayed from 6 dwts. to 8 dwts. of gold, and the 
residues, after treatment, from 1 dwt. to 2 dwts. per ton. The 
average actual extraction was about 74 per cent. 

The solutions, after leaving the precipitation boxes, still con- 
tained gold, the strong solution showing by assay 4 dwts. 8 grs., 
and the weak solution 10 grs. per ton of solution. On the aver- 
age, the strong solutions carried from 4 dwts. to 5 dwts., and the 
weak from 10 grs. to 1 dwt. of gold per ton of solution. 

From November 1894 to May 1895, the Metropolitan Com- 
pany treated 26,900 tons of tailings for 4845 ozs. of gold, at a cost 
of 2s. 8d. per ton. At the May Consolidated the working expenses 
amounted to about 2s. 4d. per ton, excluding the royalty for the 
use of the process, which amounts to 3 per cent. The extraction 
amounted to over 80 per cent, of the original assay value. 

Details of the Treatment. — The time occupied in leaching 
and washing, together with the quantity of the solutions, are given 
in the following tabulated statement : — 

Alkaline wash, 10 tons, 

Strong cyanide solution, 70 tons, 0*05 to 0*08 per cent. 
KCy, applied in 14 separate portions of 5 tons each, 

Weak cyanide solution, 21 tons, 0*01 per cent. KCy, 
applied in 3 portions of 7 tons each, 

Water washes, total 1 1 tons, pumping dry and dis- 
charging, ........ 







The working of this process gives rise to the production of a 
number of valuable commercial bye-products, including copper, 
lead, litharge, and paint. 



The Diehl Process. — This is an adaptation of the Sulman- 
Teed process. It embraces the following essential stages *: — 

1. Crushing and sliming the ore. 

2. Treating the slimes in agitators with a solution of potas- 

sium cyanide in combination with cyanogen bromide. 

3. Filter-pressing the sludge. 

4. Precipitating the gold from solution by means of zinc. 
According to the character of the gold and its associates, 

amalgamation and concentration can be added to the process. 

The most direct advantage of this process is, that the ore is 
treated in a " raw " condition. 

At Haunan Star mill, where the process was first introduced, the 
ore is dry-crushed in two No. 5 Krupp ball-mills provided with 
30-mesh screens. The crushed ore is elevated, mixed with water, 
and classified into sands and slimes. The sands are conducted to 
copper-plates on which from 10 to 15 per cent of the gold is 
saved. From the plates the pulp travels by gravitation to the 
fine-milling department, where it is ground into slimes in a large 
Krupp tube or flint mill, which is an 18 ft. long steel cylinder, 4 
ft. in diameter, charged with 4 tons of flint balls. The sand is fed 
into one end and issues at the other of such a fineness that the 
whole product will pass through a sieve with 200 mesh per lineal 

From the grinding mill the pulp is led to settling vats, from 
which the surplus water is returned to the mixing machine. The 
thickened pulp is now led into agitators, where it is treated with 
a solution of cyanide of potassium and bromide of potassium. 

When the agitator has received its full charge, a strong solution 
of potassium cyanide is added. For slimes containing 1 to 3 oz. 
of gold per ton, we have found it sufficient to add so much cyanide 
that there will be 4*4 lbs. KCy per ton of dry material. After the 

* Knutsen, Trans. Inst. Mm. and Met. , London, 1902. 



sludge has been agitated for 1 to 1£ hours, the solution of 
bromide of cyanogen is allowed to flow in, the quantity added 
being 1*1 lb. per ton of dry material. The agitator is kept going 
for 24 hours from the time the KCy solution was charged into it. 
In case the sludge should contain more than 3 oz. per ton, it may 
be advisable to add, after 6 to 8 hours' agitation, a further quan 
tity of KCy and BrCy, to ensure a good extraction. On the other 
hand, if the sludge contains less gold than 1 oz. per ton, the 
quantity of KCy and BrCy can be considerably reduced. 

About 2 hours before the agitator is ready to discharge to the 
filter-press, lime is added to the sludge in quantity varying from 
1 lb. to 4 lbs. per ton of dry slimes. In the most cases, I think, 3 
to 4 lbs. is used. A cleaner precipitate is thereby obtained in the 
zinc boxes. 

After agitation, the pulp is pressed in Dehne's filter-presses, in 
which the dissolved gold is washed out. After washing, the dry 
cakes of slime are thrown out on the dump. 

The gold in the solutions is precipitated with zinc turnings in 
the ordinary way. It is claimed that an extraction of over 93 per 
cent, can be obtained. The consumption of water is very small in 
this process, which is an important factor on a goldfield, where 
water is so scarce. All the available water is salt or brackish, 
which doubtless tends to retard close concentration. 

Mr. Feldtmann gives the following summary of the costs for the 
month of July 1901, when 2210 tons of ore were treated at 
Hannan's Brown Hill, at Kalgoorlie, by this process : — 


Milling, . 
Concentration, . 
Treatment of concentrates, 

Per ton. 











24 0-88 

The Schilz Process. — This process is based on the addition 
of barium peroxide, Ba0 2 , to the ordinary cyanide solution, which 
it is claimed under certain conditions parts with one-half of its 
oxygeu. It is further claimed that the normal oxide of barium, 
which is left after the treatment, performs further functions — 
inasmuch as it decomposes the sulphate of iron in the solution, 
forming sulphate of barium and oxide of iron, both insoluble ; it 
removes sulphocyanides, and it dispenses with the use of lime and 


answers the purpose of the same, with other minor advantages in 
addition. On the other hand, the process requires much longer 
time to effect a satisfactory extraction than in ordinary cyaniding, 
thus necessitating a considerable increase in vat capacity. 

Herr Schilz states that to ensure success the peroxide must be 
well sprinkled with the tailings whilst the vat is being filled, so 
that a good mixture may be obtained. A rather stronger cyanide 
solutiou than usual is run in, and the whole allowed to stand 
undisturbed for at least three days. The longer the time the 
better is the extraction, and in the case of concentrates it should 
be at least a week. After running off the gold solution the residue 
is washed with four weak cyanide solutions, the first of which 
should remain six hours. Then, again, a strong solution should 
be applied, followed by weak solutions. 

The quantity of peroxide depends less on the percentage of 
gold than on the length of time during which the solution remains 
in contact with the charge; and, further, the more pyrites pre- 
sent the greater the consumption of peroxide. The vat should 
not be filled to the brim with the tailings, as the mass swells up 
by the evolution of gas, sometimes as much as a foot. Thus the 
liberated oxygen should remain undisturbed as long as possible 
in contact with the charge, the greatest solvent action of the 
cyanide being after a few days. The quantity of peroxide 
required varies with the material under treatment. 

In actual practice the quantity varies from J lb. to 1 lb. per 
ton of ore treated. Delivered on the Rand, the cost of barium 
peroxide is <£50 per ton. 

A working trial of the process at the City and Suburban mine, 
according to the report of the manager, showed a gain of about 
13 per cent, iu extraction at a small additional cost — that is, a 
rise from 76 to 89 per cent. The treatment of the pyrites 
concentrates gave also satisfactory results. 

The inventor makes the following claims for his process : — 

1. Ba0 2 decomposes in contact with moistened tailings into 
BaO and 0, and supplies the cyanide of potassium with oxygen 
required for the dissolution of gold. This decomposition is 
performed in the cold. 

2. Ba0 2 supplies the oxygen which is required for the 
decomposition and dissolution of the reduced pyrites, and thus 
assists in liberating the enclosed gold. 

3. Ba0 2 in decomposing supplies BaO, which, being a strong 
basis, exerts a purifying and clearing influence upon the cyanide 
solutions, thus increasing their solving power. 

4. Ba0 2 renders superfluous the use of lime, the now existing 
use of which has in its train so many injurious secondary effects. 


The Park-Whitaker Cyanide Process.— This process is 

intended for the treatment of cupriferous ores and concentrates 
which cannot be treated successfully by the ordinary cyanide 
processes, on account of the solubility of copper ores in cyanide 

In this process the ore is subjected to a chloridizing roasting, 
after which the soluble copper chlorides are removed by leaching 
with water. An alkaline wash is then applied, and the gold and 
silver extracted with a dilute solution of cyanide. 

During the roasting the silver sulphides present are converted 
into chloride, which is readily dissolved by cyanide. The dis- 
solved copper is recovered by passing the solutions through iron 
turnings or scrap-iron. 

Experiments on a working scale were made by the author on a 
parcel of ore from the Monowai mine, N.Z., with most successful 
results, and preparations are now being made for more extensive 

The Pneumatic Cyanide Process.— In this process the 

dissolution of the gold is accelerated partly by jets of compressed 
air, which cause a continuous and gentle agitation of the pulp, and 
partly by the aeration of the solution caused by the passage of the 
air through the charge. 

The compressed air is conducted to the charge through a coil of 
perforated pipe placed on top of the filter-cloth. 

This process is identical with the Park-Horn process, for which 
provisional letters patent were obtained in New Zealand in 1895. 
Tests were made by Mr. George Horn and the author at Kuaotunu 
goldfield on excessively fine slimes. The dissolution of the gold 
was rapid and almost complete, but the mechanical difficulty of 
separating the gold-containing solutions from the slimes gave so 
much trouble that further attempts were abandoned. For the 
treatment of sands, ordinary percolation was found to give satis- 
factory results without artificial aeration. 

Gilmour- Young Process. — This process has been operated 
on argentiferous gold ores in Central America. It is essentially 
an amalgamation process, closely following the Washoe method of 
treatment. The addition of a caustic alkali, copper and zinc 
amalgam, and mercury to a pulp containing cyanide solution must 
promote many complicated reactions of doubtful utility. 



These are based on the dissolution of the gold with cyanide, 
and the electro-precipitations of the gold on mercury, which also 
serves to amalgamate any gold too coarse to be dissolved by 

The Riecken Process. — This process has lately been 
installed at the South Kalgoorlie mine, Kalgoorlie, and since 
December 1900 is reported to have treated 5000 tons of slimes. 
At present the capacity of the plant iR being doubled. The 
following details of the process are extracted from a description 
supplied to the Australian Mining Standard in 1901. 

In this process the pulverized and, if necessary, roasted ore is 
agitated in a vat of special construction with cyanide solution ; a 
current of electricity is passed through the resulting pulp, by 
which the gold dissolved in the cyanide solution is deposited on 
the amalgamated sides of the vat as amalgam, while at the same 
time any coarse particles of gold in the ore too large to be dissolved 
by the solvent are mechanically amalgamated. Thus, in one 
operation, in a single vat, are simultaneously performed — (1) 
amalgamation of the coarse gold ; (2) solution by cyanide of the 
fine gold ; (3) electrical precipitation of this dissolved gold. The 
pulp, being thus deprived of its precious metals, is at once dis- 
charged from the vat to residues dams, and the treatment is 

The apparatus in which this operation is performed may be 
regarded as a huge electro-depositing cell, which is an iron vat 
with vertical ends, inclined sides, and a rounded bottom. Its 
dimensions are 13 ft. 4 in. long, 8 ft. 3 in. wide at the top, and 11 
ft. deep. It holds a charge of 18 tons. Agitation of the pulp is 
effected by paddles attached to a horizontal shaft working through 
stuffing boxes. The paddles are 2 ft. 8 in. long, 5 in. wide, and 
spaced 12 in. from centres. The ends of paddles reach about 2 in. 
from the bottom of vat. 

The sides and bottom of vat are lined with movable amalga- 
mated copper-plates, T Y in. thick, which form the cathodes or 
negative poles. 

When in operation a thin sheet of mercury constantly descends 
over the copper-plates from a \ in. pipe perforated with ^ in. holes, 
spaced 6 in. centres. A slow reciprocating motion is imparted to 
these perforated iron pipes, causing them to swing through a chord 
of 6 in. about twenty times per minute. The mercury finds its 
way to the bottom of the vat, whence it is drawn off by a trap and 
elevated by an air-lift to the upper receptacle, whence it again 
flows into the vat, thus ensuring a continuous circulation. 


The anodes are bars of iron 1 in. thick by 3 in. wide, suspended 
from two girders and curved so as to be parallel to the cathodes. 
There are twelve cathodes in the vat. 

The agitator running about twelve revolutions per minute is 
found sufficient to keep the pulp from settling. The salt in the 
local water renders the conductivity of the pulp excellent, so that 
a current of very low potential is quite sufficient to effect the 
precipitation of the dissolved gold. The current required for 
each vat of 18 tons is about 250 amperes of a potential of 2-5 
volts, equal to about six-sevenths of an electrical horse-power, the 
cost of which per ton is merely nominal. 

The ore is first pulverized until all the gold is liberated from its 
matrix. If necessary the ore is roasted. In order to obtain good 
results it is merely necessary that the gold should be amalgam- 
able, or in such a fine state as will permit of dissolution by 
cyanide solutions. Necessarily each particular ore requires a 
certain specific preliminary preparation, determinable by experi- 
ment, to obtain the best results. 

The pulp, containing approximately equal weights of ore and 
solution, is discharged into the vat previously described, and 
which is called the " electro vat." The agitator is set in motion, 
the electric current from the dynamo is started, and the flow of 
quicksilver is maintained by means of the air-jet, as already 
described. The solution is made up of potassium cyanide to the 
average strength of 0*075 per cent. With a high-grade ore of 
2 oz. or over the above-described operation is continued for about 
eighteen hours, while with low-grade slimes, carrying but 3 dwt. 
or 4 dwt., it be may be complete in about six hours. After a 
proper time has elapsed the discharge-valve at the bottom of the 
vat is opened and the pulp is allowed to flow to the residues dam ; 
the vat is refilled as before, and the operation indefinitely repeated. 
If water is scarce, the supernatant liquor, after the pulp has 
settled in the dam, may be pumped back into the mill to make 
pulp for succeeding charges. 

The clean-up is effected by withdrawing one copper plate at 
the time, and replacing it with a spare one. The amalgam pro- 
duced is very fine, and contains about 27 per cent, of fine gold. In 
this process sodium is added to the mercury automatically to 
keep it active and bright. 

The Kiecken process dispenses with percolation, filter-pressing, 
or decantation for the removal of the dissolved gold from the ore, 
and entirely does away with zinc-precipitation and clean-up, the 
complete treatment of the ore, after its mechanical preparation, 
being effected in one operation in one vat, in the minimum of 
time, the gold being recovered as amalgam, requiring only retort- 


ing and melting. The future applications of this process will be 
watched with much interest by metallurgists. 

The Keith Electro-Cyanide Process. —The Keith electro- 
cyanide process is the invention of Dr. Keith, an American elec- 
trician. The process consists of two parts. First, the process for 
dissolving the gold out of the crushed ore ; and second, the re- 
covery of the gold from the solution. Dr. Keith's improvement 
in the dissolving process consists in adding to the solution of 
potassium cyanide a certain amount of cyanide of mercury. In 
practice, he states that he finds the best results are obtained when 
the solvent contains 0*05 per cent, of potassium cyanide and 0*025 
per cent, cyanide of mercury. This mixture of cyanides, it is 
claimed, operates very much faster than the simple potassium 

The process for the recovery of the gold from the solution is an 
electrolytic one. The gold and the mercury are deposited together 
upon amalgamated copper-plates. The amalgam so deposited is 
scraped off and the gold recovered by distilling off the mercury in 
the usual way. The anode is not allowed to dip into the cyanide 
solution, but is placed in a separate compartment and surrounded 
with a solution of an alkaline salt, so that the cyanide does not 
become decomposed. The electro-motive force of the current need 
not be more than half a volt. 



All cyanides are deadly poisons ; but the aqueous solutions used 
in practice are so dilute that there is little or no danger from the 
prussic acid evolved from them if the buildings are properly ven- 

Acids react on cyanides, liberating prussic acid gas, which causes 
almost instant death when inhaled in a pure state. When diluted 
with air, it causes faintness, dizziness, and a depressing frontal 

Even very dilute solutions of cyanide are poisonous when taken 
internally, and, when they come in contact with the skin, produce, 
in some persons, an eruption of painful red boils. In cases where 
the hands and arms must be brought into contact with the solu- 
tion, rubber gloves, reaching over the elbows, should be provided 
for the workmen. Kaffir workmen are said to suffer no incon- 
venience whatever from the contact of their skin with cyanide 

Considering the extensive use of cyanide, the number of fatal 
accidents is remarkably small. Up to the present time only a few 
fatal cases have been recorded. 

In a cyanide plant, poisoning may be apprehended from two 
principal sources, namely : — 

ia.) From hydrocyanic acid liberated in vat-house. 
b.) From poisonous gases liberated during acid treatment of 


In South Africa, Australia, and countries where the vats are not 
covered by roofs, danger from prussic acid vapours, liberated by 
the action of mineral acids or atmospherio carbonic acid, is prac- 
tically unknown. Even where the vats are enclosed in a shed, the 
risk can be reduced to a minimum by the free circulation of fresh 

The author has observed that the presence of HCy vapours is 


always more noticeable in agitation than in percolation plants, the 
obvious reason being that agitation is generally adopted for the 
treatment of pyritic ores and concentrates, while the strength of 
the solutions used is nearly always high. 

In cases of cyanide poisoning by inhaling the fumes of hydro- 
cyanic acid, a German chemist recommends the use of hydro- 
gen peroxide, H 2 2 , which forms with HCy the harmless com- 
pound oxamide, the reaction being represented by the following 
equation : — 

2HCN + H 2 2 = gS|- 

This is said to be the most reliable and satisfactory remedy 
known at the present time. 

In a case of HGy poisoning at the Crown Deep, the effect of 
the gas was immediate, one of the workmen falling as if dead. 
The same effect was observed at the N.Z. Crown mines, where a 
foreman fell into a cyanide vat without previous warning and died 

Danger during Acid Treatment of Slimes.— The slimes 

generally contain a small proportion of insoluble cyanide salts, 
which yield hydrocyanic acid when the sulphuric acid is poured 
on them. To guard against this danger, repirators should be 
used by the workmen who have to stoop over the dissolving tubs, 

In cases of poisoning, subcutaneous injections of H 2 2 are said 
by Mr. T. L Carter of the Crown Deep mine to enable the patient 
to soon come to. 

In ores containing arsenic, most of which are more or less soluble 
in cyanide, there is a danger of arsenic being precipitated with the 
gold in the zinc boxes. During the acid treatment of such slimes, 
the deadly poison arseniuretted hydrogen would be liberated by 
the action of the sulphuric acid on the zinc. 

The symptoms observed in the case of the North Pole Com- 
pany's mill, where the superintendent, and foreman both died, 
while many others were affected, were first nausea, then extreme 
langour, with pains in the legs, followed by discoloration of the 
skin in patches assuming the hue of tan ; the whites of the eyes 
became yellow as in jaundice ; finally, the passing of blood instead 
of urine to such an extent that the fluid coagulated jn a few hours, 
the patients apparently dying from internal mortification. 

The arsenic being inhaled, passes from the lungs through the 
whole system, and rapidly attacks the tissues of the body, pre- 
cluding any relief by means of antidotes. 

Where the acid treatment is used, the zinc for precipitation 
purposes should be free from arsenic ; and in all cases the dissolu- 


tion of the zinc should be conducted in a special chamber or cup- 
board connected with a chimney having a good draught. 

In cases of internal poisoning, vomiting should be induced at 
once by emetics or physical means. 

Freshly precipitated carbonate of iron, obtained by mixing equal 
quantities of sodium carbonate and ferrous sulphate, is recom- 
mended for internal use. 

It was lately reported in the press that Johann Antal, a Hun- 
garian toxicologist, had found that a solution of cobalt nitrate was 
a perfect antidote for prussic acid poisoning. Eecent investiga- 
tions, however, have shown that cobalt salts exert a toxic action 
on animals when injected subcutaneously, finally leading to death ; 
and for this reason nitrate of cobalt cannot be recommended for 
human subjects. 

Cyanide Sores. — A percentage of the workmen engaged in 
the "clean-up" of the zinc extractors are affected with painful 
sores in those parts of their hands or arms which come in contact 
with cyanide solutions. Why some men should be exempt and 
some afflicted in this way is not very clear, but it is probably 
connected with constitutional causes. 

Writing on this subject, Mr. A. Watt supplies the following 
instructive notes*: — 

" These painful affections may arise from two principal causes : 
first, from dipping the hands or arms into cyanide baths to recover 
articles which have dropped into them — a very common practice 
and much to be condemned; and second, from the accidental 
contact of the fingers or other parts of the hand, on which a recent 
cut or scratch has been inflicted, with cyanide solutions. In the 
former case, independent of the constitutional mischief which may 
arise from the absorption by the skin of the cyanide salts, the 
caustic liquid acts very freely upon the delicate tissue of the skin, 
but more especially upon the parts under the finger nails. We 
have known instances in which purulent matter has formed under 
the nails of both hands from this cause, necessitating the use of the 
lancet and poulticing. Again, when cyanide solutions come in 
contact with recent wounds — even very slight cuts or abrasions of 
the skin — a troublesome and exceedingly painful sore is sure to 
result, unless the part be at once soaked in warm water ; indeed 
it is a very good plan, after rinsing the part in cold water, to give 
it a momentary dip in a weak acid pickle, then soak it for a few 
moments in warm water, and after wiping the part dry with a 
clean rag or towel, apply a drop of olive oil and cover up with a 
strip of thin sheet of gutta-percha." 

* Watt, Electro-Deposition, p. 611. 


Provision of Remedies. — In order to minimize the danger 
attending cyanide poisoning, the necessary antidotes should be 
provided in every cyanide plant, and kept in a closed cabinet with 
a glass front, placed in some conspicuous and easily accessible 
part of the works, known to all the workmen. 

The cabinet should have the words Antidotes for Cyanide 
Poisoning marked in clear letters near the top, and written or 
printed instructions how to apply the remedies pasted inside the 

The cabinet should contain the following articles : — 

1. A sterilized glass flask marked A, filled with a 3 per cent, 
solution of hydrogen peroxide. The neck of the flask should be 
drawn to a fine point and sealed in a blow-pipe flame. 

2. A sterilized glass flask marked B, filled with a 30 per cent, 
solution of hydrogen peroxide, sealed with flame. 

3. A hypodermic syringe, made of glass. 

4. A stomach tube and funnel. 

5. A small conical-shaped medicine glass. 

6. A small triangular file. 

7. A small pair of pinchers. 


Absorption of cyanide by vats, 6. 
Acid ores, preliminary washes for, 69. 

slimes, smelting of, 108. 

tailings, 76. 

treatment of slimes, 110, 111, 

Africa, acid treatment of slimes in, 

111, 113. 

wet crushing in, 125. 

Agitation leaching, 81, 94, 96. 

actual extraction by, 97. 

Agitators, 82, 86, 91, 96, 97, 138, 

151, 177. 
Air, compressed, use of, 180. 

compressors, 65, 83. 

lifts, 65, 66. 

for slimes, 63. 

practice at Ealgoorlie, 64. 

pumps, 55. 

Alkali, protective, 166. 

estimation of, in solutions, 

Alkaline cyanides, tests for, 34. 

sulphides, influence of, 14. 

wash, 75. 

Amalgamation processes, 180, 181. 

See also Pan-amalgamation. 
America, cost of Zn precipitation in, 


dry crushing in, 1 25. 

sulphuric acid refining in, 110, 

Analysis of solutions, 22. 
Analytical methods, 33. 

Crosse's, 41, 43. 

Feldtman's, 35. 

Green's, 36. 

Virgoe's, 33. 

Anodes, 173, 181. 

Antidotes for cyanide poisoning, 185. 

Antimonite, influence of, 11. 

Antimony, influence of, 13. 
Appliances used for leaching, 48. 
Assay of cyanide solutions, 31. 

tables, 46. 

Associated and Westralia Company, 

air-lift at, 65. 
Athabasca Mine, practice at, 167. 
Azurite, influence of, 10. 

Banket, 121. 

Barium peroxide, use of, 178. 

Base metals, estimation of, in solu- 
tions, 43. 

Bonanza Mine, Johannesburg, Tave- 
ner process at, 115. 

British Columbia, practice in, 167. 

Bromo-cyanogen, use of, 178. 

Bucket- wheels, 63. 

Butters' discharging doors, 57. 

distributes, 73, 155, 165. 

Calculating percentage extraction, 

California, practice in, 161. 
Camp Bird Mines, acid treatment at, 

practice at, 15, 102, 103, 

Carbonic oxide, influence of, 7. 
Cassel Gold Extraction Company, 11, 


practice of the, 96. 

tell-tales at, 56. 

Cathodes, 173. 

Caustic lime, use of, 75. 

soda, use of, 75. 

Centrifugal pumps, 152. 
Chalcopyrite, influence of, 12. 

in sulphide ores, 93. 

Channels, formation of, 83. 



Charcoal, influence of, 11. 

precipitation, 12, 120, 154. 

Cinnabar, occurrence of gold with, in 

Utah, 162. 
City and Suburban Mine, direct filling 

at, 74. 

— : practice at, 179. 

Clarifying presses, 152. 
Clean-up, 106. 

in electrical processes, 175. 

Cobalt, influence of, 15. 
Colorado : see Camp Bird Mines. 
Combined leaching and agitation, 82. 
Compressed air, use of, 180. 
Concentrates, filling vats with, 72. 
treatment of, 89, 93. 

value of, 94. 

Concentration of solutions, 86. 

Constants, some useful, 47. 

Construction of vats, 52. 

Consumption of cyanide, 5. 

of zinc in precipitation, 101. 

Copper, influence of, 12, 94. 

on zinc precipitation, 102, 

103, 104. 

use of, 157. 

ores, influence of, 9. 

Cost of plant, 67. 

of tailings treatment, 126. 

Covelline, influence of, 12. 

Cripple Creek, vats at, 53, 85, 86. 

Crown Mines, Earangahake, 96, 128. 

Crown Reef, cost of treating concen- 
trates at, 93. 

discharging vats at, 59. 

filling vats at, 74. 

vats used at, 53. 

Crude cyanide, strength of, 30. 

Cupriferous ores, treatment of, 180. 

Current required for precipitation, 

Cyanide, consumption of, 34. 

extraction, appliances for, 48. 

loss of, 6, 133. 

plant, at Thames School of 

Mines, N.Z., 97. 

poisoning, 184. 

- solutions, analysis of, 33, 35. 
assay of, 31. 

sores caused by, 186. 

CyanidinginN.Z., 127. 

in presses, 88. 

Cyanogen bromide, 177. 

Dead roasting, tests for, 89. 
Decantation, 84, 161. 

Dehne filter-press, 88, 89, 145, 146, 

Diehl process, 89, 145, 149, 177. 
Dilution of solutions, 29. 
Dioptase, influence of, 13. 
Direct filling, 72, 74, 123. 
Discharge doors, 57. 

Butters', 57. 

Feldtman's, 59. 

Irvine's, 58. 

Roche's, 59. 

of leached residues, 56. 

vats, 74. 

Dissolving tanks, 50. 
Distributors, 73, 130, 155, 165. 

automatic, 73. 

Butters', 73, 155, 165. 

Doors, discharging, 57. 
Double cyanides, estimation of, in 
solutions, 36. 

treatment, 71. 

Dry crushing, 80. 

American practice, 125. 

cost of plant for, 67. 

filling vats after, 

number of vats for, 66. 

v. wet cutting in N.Z., 127. 

ore vats, 66. 

Drying acid-refined slimes, 112. 

Electrical precipitation, 171, 173, 


advantages of, 175. 

boxes for, 175. 

clean-up after, 175. 

current required for, 173. 

Electro-chemical processes, 84, 181. 
Electrolysis of solutions, 172. 
Elnsner's reaction, 4, 5, 132. 
Estimation of strength of solutions, 

Experiments, agitation leaching, 99. 

on strength of solutions, lb. 

Extraction, 69. 

by agitation, 97. 

rate of, 98. 

Extractor boxes, 125. 
Extractors, zinc, 60, 120, 125. 

Ferro-cyanides, estimation of, 36. 
Filling, direct, 72, 74. 

intermediate, 62. 

tailings, 72. 

vats, 69, 71. 

Filter frames, 54, 55. 
presses, 63, 88, 152. 



Filter-presses, methods, 84. 

practice, N.Z., 87, 151. 

vats, 63. 

Fineness required, 81, 84. 
"Float" gold, 80. 
Fluxes, for lead smelting, 117. 
Franklinite in gold concentrates, 93. 
Free milling ores, 17. 

sulphur, influence of, 13. 

Freely percolating tailings, vats for, 

Furnace for melting slimes, 109, 117. 
Fusion of gold slimes, 109. 

Galena, influence of, 13. 

in sulphide ores, 93. 

Gilmour- Young process, 180. 

Gold, combination of, with sulphur, 

precipitation of, by zinc, results 

of, 106. 

slimes, 107. 

— precipitation of, by zinc, 


smelting of, 109. 

Golden Gate Cyanide Works, plant 

at, 163. 

Horseshoe, practice at, 147. 

Grain and gram table, 46. 

Great Boulder Proprietary, practice 

at, 145. 
Mercury Cyanide Works, 75. 

Han nan Star Mine, Diehl process at, 

Hauraki Goldfields, 127, 133. 

consumption of cyanide at, 

132, 133. 

Homestake Company, Dakota, 120, 
154, 164. 

Hydrocyanic acid, estimation of, 35. 

poisoning by, 184. 

Hydrogen, evolution of, in acid refin- 
ing, 111. 

in leaching, 158. 

in zinc precipitation, 102, 


India, practice in, 169. 

sun-drying in, 92. 

Intermediate filling, 73. 

Iodine solution, standardising, 26. 

Iron pyrites, influence of, 7, 13. 

occurrence of, 93. 

salts, influence of, 8, 9. 

Irvine's discharging doors, 58. 

Johannesburg, cost of plant at, 67. 

electrical precipitation at, 174. 

practice in, 73. 

Tavener process in, 115. 

Jumpers Deep, new plant at, 122. 

Kalooorlie, air-lift at, 64. 

practice in, 87, 89, 144, 146, 

Eapai Vermont Works, 12, 144. 
Earangahake, 10, 75. 
Kauri Gold Estates, ores from, 128. 
Eeith electro-chemical process, 183. 
Eiln-dried ores, behaviour of, 7, 12. 
Eomati Gold Mine, ores from, 15. 
Eoppel patent tank doors, 58. 
Euaotunu, practice at, 12, 75, 92. 

104, 180. 

Laboratory routine, 17. 

Lake View Consols, ore from, 87, 148. 

Langlaagte Estate Co.'s vats, 53. 

practice at, 57, 72, 109. 

Leached residues, discharge of, 56. 
Leaching, 70, 76. 

by agitation, 81, 82, 94, 96. 

cost of, 67. 

vats for, 48, 51. 

Lead, influence of, 13. 
acetate, use of, in zinc precipita- 
tion, 103. 

couple for zinc precipitation, 102. 

precipitation, 175. 

smelting, 115. 

cost of, 119. 

fluxes for, 116. 

Loss of cyanide, 6, 133. 

of gold, on cupellation, 119. 

Luipaards Wei Estate, practice at, 

Lydenburg, precipitation practice at, 


M c Arthur- Forrest process, 1. 
Main Reef, filter frames at, 55. 
Malachite, influence of, 10, 13. 
Manganese dioxide, influence of, 15. 

in acid refining, 111. 

Marlborough, N.Z., experiments at, 

Martin press, 88. 
Masonry vats, 53. 
Melting gold slimes, 109. 
Mercur, practice at, 162. 
Mercuric chloride solution, 25. 



Mercury couple, in zinc precipitation, 


influence of, 11. 

use of, 157, 181. 

in India, 169. 

Metallic sulphides, influence of, 12. 
Methods of filling vats, 72. 
Mineral acids, influence of, 7. 
Moanataiari, discharge doors at, 60. 

practice at, 58, 89, 134. 

vats used at, 52, 53. 

Monowai, experiments on ores from, 

81, 83. 

practice at, 102, 133, 134, 180. 

sulphides, treatment of, at, 98. 

Montana, practice at, 161. 

zinc precipitation at, 120. 

Montejus, 146, 151, 152. 

Mount Malcolm Proprietary, air-lift 

at, 63. 
Mysore Works, practice at, 167. 

Natural settlement, 85. 
Nevada, practice in, 161. 

sun-drying in, 92. 

zinc precipitation in, 161. 

New Eleinfontein, tailings practice 
at, 123. 

Mexico, practice in, 161. 

South Wales, practice, 151. 

Zealand, concentrate treatment 

in, 94. 

practice in, 56, 127. 

Nickel, influence of, 15. 

Nitre, use of, in smelting, 107, 170. 

Ohinemuki, occurrence of gold at, 

practice at, 134. 

Order of operations, 71. 
Ores, kiln-dried, 7, 12. 

testing of, 17. 

Oxidising agents, 69. 
Oxygen, influence of, 15. 
in working solutions, 41. 

Pan amalgamation, 127. 
Park-Horn process, 180. 

Whitaker process, 180. 

Percentage extraction, 19, 20. 
Percolation, leaching concentrates by, 

plant at Crown Mines, 96. 

vats, 51. 

Phenolphthalein indicator, 37. 
Plant, cost of, 67. 

Pneumatic cyanide process, 180. 
Poisoning, antidotes for, 184. 
Precipitation boxes, 175. 

by charcoal, 12, 120. 

by zinc, 100, 120. 

tanks, 61. 

Preliminary wash, 69, 75. 
Pressing cakes, 151. 
Protective alkali, 166. 

- estimation of, 44. 

Prussian blue, formation of, 9. 

— recovery of cyanide from, 

Prussic acid, liberation of, 184. 
Pumps for solutions, 63. 
Pyrites, influence of, 13, 94. 
Pyritic concentrates, 93. 

ores, 7. 

loss of cyanide by, 75. 

testing for, 19. 

treatment of, 70. 

tailings, zinc precipitation of, 

Pyrolusite, influence of, 15. 

Rand Central Reduction Company's 

plant, 84. 
Rate of extraction, 18. 

of solution, 14. 

Reactions involved, 4, 8, 9. 
Remedies for poisoning, 186. 
Residues, discharge of, 56. 
Riecken process, 89, 145, 150, 181., 
Roasting before cyaniding, 89. 

gold slimes, 107. 

Robinson Works, 85. 

Roche discharging doors, 59, 137. 

Sands, treatment of, 91. 
Schilz process, 179. 
Separators, 86, 87. 
Sequence of operations, 71. 
Settlement, 86. 
Siemens-Halske process, 171. 
Silver, influence of, 133. 

nitrate solution, 22. 

Simmer and Jack, vats used at, 52. 
Slags from acid refining, 113. 

from lead smelting, 119. 

from smelting slimes, 109. 

Slime cake pressing, 140. 
— mixers, 151. 
Slimes-arrester, 167. 

refining by sulphuric acid, 110. 

smelting of, 107. 

sun-drying of, 92. 



Slimes, treatment of, 80, 83, 92, 107, 

110, 140. 

at Waihi, 140. 

by Tavener process, 115. 

Slimy sands, leaching, 79, 161. 

tailings, 66, 70. 

Smelting acid slimes, 108. 

gold slimes, 107. 

Soda, use of, in acid smelting ol 

slimes, 108. 
Solution pumps, 63. 

rate of, 4. 

vats, 50. 

Solutions, bulk of, 122, 124. 

dilution of, 29. 

for testing, 19. 

titration of, 22. 

weak, use of, 70. 

Sores, from cyanide poisoning, 186. 
South Dakota, practice in, 164. 
Spitzkasten, 72, 86, 140, 147. 
Spitzluten, 86, 138, 139. 
Stamp batteries, 80. 


Steel tanks for zinc precipitation, 


vats, 54, 67. 

Stibnite, influence of, 13. 
Strength of solutions contrasted, 5. 
Strong solution leaching, 76. 
Sulman-Teed process, 177. 
Sulphide ores, 77. 

— extraction of, by agitation, 


treatment of, 149. 

Sulphides, alkaline, influence of, 14. 

roasting, 148. 

Sulpho-cyanides, action of, 15. 

estimation of, 36. 

Sulpho-telluride ores, 144. 
Sulphur, free, influence of, 13. 
Sulphuric acid refining, 110. 

cost of, 113. 

Sumps, 60. 

washes, 77. 

Sun-drying, 84, 92, 161. 

Tables for assay solutions, 46. 
Tailings, filling vats with, 72. 

test solutions for, 19. 

treatment, at Waihi, 142. 

value of, on the Rand, 122. 

vats for, 52. 

Talisman Mine, Ohinemuri, practice 

at, 134. 
Tank, dissolving, 50. 

Tavener process, 115. 

cost of, 119. 

Thames Experimental Works, 79, 97, 

133, 170. 
Goldfield, concentration plant 

at, 96. 

School of Mines plant, 97. 

Tell-tales, 63. 

Telluride ores, treatment of, 144. 

practice, at Cripple Creek, 160. 

Testing crude cyanide, 30. 

for acidity, 34. 

solutions, 22. 

strength of solutions, 18. 

Tests for consumption of cyanide, 34. 

Threshis' method, 158. 

Titration of standard solutions, 23, 

Treatment of slimes, 80. 
of tailings, on the Rand, 123. 

Upward filtration, 83. 
Useful constants, 47. 
Utah, practice in, 162. 
zinc precipitation in, 120. 

Vacuum cylinders, 55. 
Vats for solutions, 50. 

construction of, 52. 

iron, 51. 

wooden, 51, 67. 

filling, 72, 81. 

leaching and percolation, 51. 

Victoria, charcoal precipitation in, 

Wad, influence of, 15. 

Waihi, extraction practice at, 62. 

filter frames at, 55. 

general practice at, 11, 56, 84, 

86, 99, 135. 

vats used at, 51, 53, 57. 

Waikino, new plant at, 59, 62, 122. 

practice at, 142. 

Waitekauri, practice at, 15, 131, 144. 
Washes, 78. 

preliminary, 69. 

water, 78. 

weak cyanide, 78. 

Washoe process, 180. 

Weak solutions, precipitation from, 

Weights and measures, 47. 



Western Australia, practice in, 144, 

sulpho-telluride ores, treat- 
ment, 144. 

Westralia Gold Mining Company, 
electrical precipitation at, 174. 

Wet crushing, 80, 90, 120, 128. 


on the Rand, 120. 

Witwatersrand, losses on the, 6. 

electrical precipitation on the, 


practice on the, 56, 70, 72, 75, 

121, 171. 
Wooden vats, 51. 

cost of, 67. 

Woodstock Gold Mining Company, 

agitation practice at, 94, 
drying ores at, 75. 

Yellow Creek, Kirk, plant at, 165. 

Zinc boxes, 100. 

consumption of, 101. 

estimation of, in solutions, 38, 

— extraction, 60, 120. 

influeuce of, 18. 

precipitation, 100, 120, 125. 

cost of, in America, 120. 

extractor boxes, 125. 

influence of copper on, 103, 

influence of lead on, 102, 


refining slimes by sulphuric 
acid, 110, 113. 

slime treatment by lead 

smelting, 107, 115. 

Tavener process, 115. 

use of mercury in, 103. 

turnings, treatment of, 

103, 105. 





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Beautifully Illustrated. With a Frontispiece in Colours, and Numerous 
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The Spacious Air. — The Open Fields and Downs. — In the Hedgerows. — On 
Open Heath and Moor.— On the Mountains.— Amongst the Evergreens.— 
Copse and Woodland.— By Stream and Pool. — The Sandy Wastes and Mud- 
flats.— Sea-laved Rocks.- Birds of the Cities.— Index. 

"Enriched with excellent illustrations. A welcome addition to all libraries." — West' 
■minster Review. 



Third Edition, Revised and Enlarged. Large Crown 8vo, with numerous 

Illustrations. 3s. 6d. 


As Illustrating the First Principles of Botany. 


Prof, of Biology, University College, Aberystwyth; Examiner in Zoology, 

UnWfcrsity of Aberdeen. 

" It would be hard to find a Text-book which would better guide the student to an accural* 
bnowledge of modern dtscoreries in Botany. . . ■ The scientific accuracy of statement, 
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the chapter on the Physiology of Flowers, an admirabls rtswni, drawn from Darwin, Herman* 
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Journal of Botany. 


With Illustrations. Crown 8vo. Cloth. 3s. 6d. 


A Simple Introduction to Real Life in the Plant-world, Based on Lotions 

originally given to Country Children. 


*»• The attention of all interested in the Scientific Training of the Young is requested to this 
osjusHTf ullt jesse and chakmino iiTTLB book. It ought to be in the hands of erery Mother 
sad Teacher throughout the land. 

" The child's attention is first secured, and then. In language simflb, tst scibhtivicauy 
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With Illustrations. Crown 8vo. Gilt, Ss. 6d. 



And other Studies from the Plant World. 
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Author of How Plant* Live and Work. 

" A, bright little introduction to the study ot Flower*:'— Journal of Botany. 
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Griffin's Standard Publications 



^PJ!^ * eeh a ni <>S> ■ Rankine, Bbownk, Jamieson, 35, 47*84 

Civil Engineering, . 
Design of Structures, . 
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Engineering Drawing, . _ _. „_ D , 
Central Electrical Stations, C H. Wordingham, 

Pbop. Rankinb, 
8. Anglin, . 
Prop. Fidler, 
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L. Andrews, 

W. H. Cole, . 

Santo Crimp, 

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Prof. A. Jamieson, 

W. F. Pettigrew, 

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Rankine, Jamieson, . 36; 34 

Bryan Donkin, 
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Of, Bryan Donkin, . 
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Electrical Pocket-Book, Munro and Jamieson, 

Electrical Price-Rook, . H. J. Dowsing, . 
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Third Edition, Revised, with an Additional Chapter on Foundation*- 
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A Ppaujtteatf Trsjatlojsj on th# Bulletin* of Brtdffos, 

By S. ANGLIN, C.E., 

If aster of Engineering, Royal Uniyersity of Ireland, late Whitworth Scholar, &c 

' Students of Engineering will find this Text-Book invaluable —AtxkiUct 

"The author has certainly succeeded in producing a thoroughly practical Text- 
Book. '-4«06r. 

''We can unhesitatingly reoommend this work not only to the Student, as the 
Thstt-Book on the subject, but also to the professional engineer as an sxcsjom: 
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Third Edition, Thoroughly Revised. Royal 8tv. With numerous 
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Being a Text-Book on the Construction of Bridges 1b 

Iron and SteeL 


Prof of Engineering, Unnrersity College, Dundee 

General Contents. — Part I.— Elementary Statics. Part II. — 
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Materials. Part IV. — The Design of Bridges in Detail. 


The new edition of Mr. Fidler's work will again occupy the same con- 
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aooorded to its predecessors. The instruction imparted is bound, simple, 
and full. The volume will be found valuable and useful alike to those who 
may wish to study only the theoretical principles enunciated, and . • - 
to others whose object and business is . . . practical." — The Engineer. 



At Press. In Large 8vo. Handsome Cloth. With Copious Plates 

and Illustrations. 

The Principles and Practice of 


By BRYS30N CUNNINGHAM, B.E., Assoc.M.Inst.CE., 

Of the Engineers' Department, Mersey Docks and Harbour Board. 


Historical and Discursive. — Dock Design.— Constructive Appliances. — 
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*«* The object of the Author has been to deal fully and comprehensively with the 
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It is primarily intended for the student ; but it is hoped that the large amount of data 
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Third Edition. In Two Parts, Published Separately. 


Engineering Drawing and Design 

Vol. I. — Practical Geometry, Plans, and Solid. 3s. 

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▲.M.INST.CB., ▲.M.IN8V.1MCH.*., 

Principal of the Batters** Polytechnic Institute, and Head of the Engineering Department 

therein ; formerly of the Engineering Departments of the Yorkshire OoTlege, 

Leeds ; and Dulwich College, London. 

With many Illustrations, specially prepared for the Work, and numerous 
Examples, for the Use of Students in Technical Schools and Colleges. 

'* A capival tixt-book, arranged on an ixcillbxt btstbm, calculated to give an intelligent 
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"The first book leads iasilt and matukallt towards the second, where the technical pupil 
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Works by BRYAN DONKIN, M.Inst.C.&, M.Inst.Meeh.E., &e 

Third Edition, Revised and Enlarged. With additional 
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A Practical Text - Book on Internal Combustion Motors 

without Boiler. 

By BRYAN DONKIN, M.Inst.C.E., M.Inst.Mech.E. 

GsiiaiiAL CoNTBMTt.— tias Engine* :— General Description— History and Develop* 
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Theory of the Gas Engine — Chemical Compoeition of Gai in Gaa Engine*— Utilisation of 
Heat — Explosion and Combustion. Oil MOtOPS :— History and Development— Various 
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awa*— Various Types: Stirling's, Ericsson's, etc., &c. 

"The iht book now published on Gsj, Oil, and Air Engines. . . . Will he of 
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and an accurate perception of the requirements of Engineers. " — Ths Engintsr. 

"We hxabtily KBCOMMBND Mr. DonMn's work. ... A monument of caveful 
labour. . . . Luminous and comprehensive." — y*w n*l »fG+s Lighting. 

" A thoroughly ksliablb and exhaustive Treatise." — Engineering. 

In Quarto, Handsome Cloth. With Numerous Plates. 25s. 



With many Tests and Experiments on different Types of 

Boilers, as to the Heating Value of Fuels, &e., with 

Analyses of Gases and Amount of Evaporation, 

and Suggestions for the Testing of Boilers. 


General Contents. — Classification of different Types of Boilers— 
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shown in Fifty Tables — Fire Grates of Various Types — Mechanical Stokers — 
Combustion of Fuel in Boilers — Transmission of Heat through Boiler Plates, 
and their Temperature — Feed Water Heaters, Superheaters, Feed Pumps, 
etc. — Smoke and its Prevention — Instruments used in Testing Boilers — 
Marine and Locomotive Boilers — Fuel Testing Stations — Discussion of the 
Trials and Conclusions — On the Choice of a Boiler, and Testing of Land, 
Marine, and Locomotive Boilers — Appendices — Bibliography — Index. 

With Plates illustrating Progress made during recent years, 
and the best Modern Practice. 

"A work 07 BxraaxvcB at fbzskmt uhiqus. Will give an answer to almost any 
question connected with the performance of boilers that It is possible to aok."*— .sTnoifMer. 

"Probably the most xxhaustivb ruurni that haa ever been collected. A pbaoxtoal 
Book by a thoroughly practical man."— iron and Coal Trade* Renew. 



Tun* Bbitiom, Rannd and Btdarfd. PoskstStu, L**th$r % 12*. UL; also Larfer Mw fat 

OJkt Use, Cloth, 12*. 6d 

Boilers, Marine and Land: 







By T. W. TRAILL, M. Inst. 0. E., F.E.RN., 

Late Engineer 8urveyor-in-Chief to the Board of Trad*. 

* # * To the Second and Thibd Editions many New Tables for Preshubs, 
up to 200 Lbs. per Square Inch have been added. 

" Thi most valuable woax on Boilers published in England."— Shipping World. 

Oontaina an Rjtobhous Quantity or Information arrranged in a rery oonvenient form. . . 
A most U8K7UL volttmb . . . supplying information to be had nowhere else."— Th$ BngMumr. 

Third Impression. Large Crown 8vo. With numerous Illustrations. 6s. 


A Handbook for Engineers and Officers in the Royal Navy 

and Mercantile Marine, Including the Management 

of the Main and Auxiliary Engines on 

Board Ship. 


Bngineer, B.N., A.M.I.C.E., Instructor in Applied Mechanics at the Royal Naval 

College, Greenwich. 

Oontmts.— General Description of Marine Machinery.— The Conditions of Service and 
Duties of Engineers of the Royal Navy:— Entry and Conditions of Service of Engineers of 
the Leading S.S. Companies.— Raising; Steam— Duties of a Steaming Watch on Engines 

and Boilers.— Shutting off Steam.— Harbour Duties and Watches Adjustments and 

Repair* of Engines.— Preservatio and 1 epairs of "Tank" Boilers.— The Hull and its 
Fittings.— Oleaningand Painting Machinery —Reciprocating Pumps, Feed Heaters, and 
Automatic Feed -Water Regulators. — Evaporators. — Steam Boats. — Electric Light 
Machinery.— Hydraulic Machinery.— Air-Compressing Pumps.— Refrigerating Machines. 
—Machinery of Destroyers.— The Management of Water-Tube Boilers.— Regulations for 
Entry of Assistant Engineers, R.N.— Questions given in Examinations for Promotion of 
Engineers, R.N.— Regulations respecting Board of Trade Examinations for Engineers, fto, 

" The contents cannot wail to bk appuciatsd."— Th* Steamship. 

" This vbbt useful book. . . . Illubtbations are of GKBAT iMPOBTANOz in a work 
•f this kind, and it is satisfactory to find that sfsclal attention has been given in this 
respect. "—Enginter*' OasetU 

In Grown 8vo, soUra, with Numerous Illustrations, [Shortly, 


An Introductory Text-Book on the Theory, Design, Construction, 
and Testing of Internal Combustion Engines without Boiler. 


By Prof. W. H. WATKINSON, Whit. Sch., M.Inst.Muoh.E., 

Glasgow and West of Scotland Technical College. 



8bookd Edition, Reyised. With numerous Plates reduced fi 
Working Drawing! and 280 Illustration! in the Text. 21s. 



A Practical Text-Book for the Use of Engine Boilden, 

Designers and Draughtsmen, Railway 

Engineers, and Students. 


With a Section on American and Continental Bnginee. 


Of Hii Majesty's Petent Oftec. 

Oontmf. — HlstoriosJ Introduction. 17U-1SSS. — Modern Looomoares : Simple. — 
Modem Loeomotrfos : Compound.- Primary Conalderagom in LoeomovWe Design.— 
Cylinders, Steam Ghetto, and ttufsng Boxes.— Pistons, Piston Bode, Orosshsads. and 
■fide Ban.— Oonneetiatand Coupling Bode.— Wheels and Axles, Axle Boxes, Horabloeka, 
end Bearing Bprtags.— Balancing .— YalTe Gear.— Slide Yelres and Valye Gear Detail*— 
framing, Bogles and Axle Tracks, Badlel Axle Bozee.— Boilers.— Smokebox. Blaet Pipe, 
firebox Fitttags.— Boiler Mountings.— Tenders.- Bsilway Brakes.— LubrioaMon.— Oon- 
rauptlon of TueL BYaporatlon and Bngiae Emolenoy.— Amerioan LooomotiTee.— Oen- 
ttnental LoeomoilTee.— Bepsirs, Banning, Inspection, and Benewale.— Three Appondloes. 

44 Likely to remain for many years the Standard Work for those wishing to learn 

u A most Interesting and Tamable addition to the bibliography of the Locomotors."— 
ReHtwav OJlcial QatstU. 

44 We recommend the book as thoroughly ibaotkial in its character, and Msarrnre A 
nAORix am ooluotxov of . . . works on LooomotiTe Bngineeriiig."~£«tf«>ejr Ifmct. 

"Tee work oovtadts all that oast bb lrabut from a book upon smoh a subjeot. It 
will at onoe rank as tmb stabdahd woes ufox ran ncronrAirr subjhot."— Rmihsaf Magaain*. 

In Large Svo. Handsome Cloth. With Plat— and Illustration*. 16s. 



Late Deputy-Manager, North- Western Railway, India. 

Con^to.— Discussion of the Term <( Light Railway*. "—English Railways, 
Rates, and Fanners. — Light Railways in Belgium, France, Italy, other 
European Countries, America and the Colonies, India, Ireland.— Road Trans- 
port ae an alternatiye.— The Light Railways Act, 1896.— The Question of 
Gauge. — Construction and Working. — Locomotives and Rolling-Stock. — light 
Railways in England, Scotland, and Wales.— Appendices and Index. 

u Mr. W. H. Oole has brought together ... a labor axouvt of yaluarlh ixtosxa- 
txon . . . hitherto practically inaooessible to the ordinary reader."— Tlmu. 

44 Will remain, for some time yet a Standard Work in everything relating to Light 
BaUwarB.''— J»^»'»Mer. 

44 The author has extended practical experienoe that makes the book moid and usefmL 
It is RXCRRDnieLT well done. "— amsin%stri n g. 

44 The whole subject is bxhacstitrlt and raAonoALLT considered. The workoambe 
cordially recommended as nmarsHSAHLB to those whose duty it is to become acquainted 
with one of the prime necessities of the immediate future."— RaUway OJkiml QmstU. 

"Tmsui could BR mo BRTTRR book of first reference on its subjeot. All olaseet of 
engineers will weloome its appearance." — Scotsman. 




Third Edition, Revised and Enlarged. With Numerous 

Illustrations. Price 8s. 6d. 





CHARLES HURST, Practical Draughtsman. 

** Oobcisb explanation* illustrated by 116 ybry cilia* diamams and drawings and 4 folding* 
plates . . . th« book fulfil* a t alxj A»Lifonctlon."—^<A<namm. 

" Mb. Huksv'b valves and yalvb-cbabivs will prove a very valuable aid, and tend te thr 
produetlon of Engines of sgibvviiio dbsi«x and bcovomioal wobkina. . . . Will b« largely 
■ought after by Students and Designers."— Mmrimt Snoinmr. 

** TJsbtul and tmokoucmlt fbaotioal. Will undoubtedly be found of «kbat talus to 
ail concerned with the design of Valve-gearing."— Moehanital World. 

44 Almost bvsby rm of valvb and its gearing is dearly set forth, and illustrated in 
such a way as to be bbadily uvdhmtood and fbaotioally apfubd by either the Engineer, 
Draughtsman, or Student. . . . 8hould prove both usbtul and valuablb to all Engineer* 
seeking for bbliablb and glbab information on the subject. Its moderate price brings it 
within the reach of %\L"—lndu*iriuan<{Iron. 

M Mr. Hubby's work is admtbably suited to the needs of the practical meohanio. . . . 
It is free from any elaborate theorettoal discussions, and the explanations of the various 
types of valve-gear are aooompanled by diagrams whioh render them basily uwdemtoob " 
— f%s Scientific A meri tmn. 

NlntB on Steam Engine Design and Construction. By Charles- 
Hurst, "Author of Valves and Valve- Gearing.'* In Paper Boards, 
8vo., Cloth Baok. Illustrated. Prioe Is. 6d. net. 

Oontbnts.— I. Steam Pipes.— II. Valves.— III. Cylinders.— IV. Air Pumps and Con - 
denBers.-V. Motion Work.— VI. Crank Shafts and Pedestals.— VIL Valve Gear. -VIII. 
Lubrication.— IX. Miscellaneous DetailB — Ikdbx. 

"▲handy Yolume whioh every practical young engineer should possess."— The Model 
Engineer . • 

JUST OUT. Strongly Bound in Super Royal 8vo. Cloth Boards. 

7s. 6d. net. 

For Calculating Wages on the Bonus or Premium Systems^ 

For Engineering, Technical and Allied Trades. 


Technical Assistant to Messrs. Bryan Donkin and Clench, Ltd., and Assistant Lecturer 
in Mechanical Engineering at the Northampton Institute, London, E.C. 

44 The adoption of this system for the payment of workmen Has created a demand for 
some handy table or series' of tables, by means of which the wages may be easily found 
without the necessity of any calculations whatever. With the object of supplying this 
need, the author has compiled the following tables, which have been in practical use- 
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additional advantage of being less liable to error, as there is practically no possibility of a 
mistake occurring. *— Extract from Preface. 



Largo 8vo, Handsome Cloth. With Illustrations, Tablet, &o. 21 1. 

Lubrication & Lubricants: 




Chemist to the Midland Railway Company, 


Midland Railway Locomotive Works' Manager, Derby. 

COHTBNT8.— I. Friction of Solids.— II. Liquid Friction or Viscosity, and Plastic 
Friction.— III. Superficial Tension.— IV. The Theory of Lubrication,— v. Lubricants, 
their Sources, Preparation, and Properties.— VI. Physical Properties and Methods of 
Examination of Lubricants.— VII. Chemical Properties and Methods of Examination 
of Lubricants. —VIII. The Systematic Testing of Lubricants by Physical and Chemical 
Methods.— IX. The Mechanical Testing of Lubricants.— X. The Design and Lubrication 
of Bearings.— XI. The Lubrication of Machinery.— Ihpkx. 

" Destined to become a classic on the subject."— Industries and Iron. 
" Contains practically ALL THAT 18 KVOWN on the subject Deserres the careful 
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Fourth Edition. VoryfuUy IUush-cUsd. Cloth, U. 64. 




Cats/ Bnginssr of tks Scottish BoUsr Insurants and Engins Inspection Company. 

Gbkkkal Contents.— I. Explosions caused (x) by Overheating of Plates— (a) By 
De fective and Overloaded Safety Valves — (3) By Corrosion, Internal or External — (4) By 
Defective Design and Construction (Unsupported Flue Tubes ; Unstrengthened Manholes ; 
Defective Staying; Strength of Rivetted Joints; Factor of Safety)— II. Construction op 
Vertical Boilers: Shells— Crown Plates and Uptake Tubes— Msn-Holee, Mud-Holes, 
and Fire-Holes — Fireboxes — Mountings — Management — Cleaning — Table of Bursting 
Pressures of Steel Boilers — Table of Kivetted Joints — Specifications and Drawings of 
Lancashire Boiler for Woiking Pressures (<*) 80 lbs. ; (i) 200 lbs. per square inch respectively. 

" A valuable companion for workmen and engineers engaged about Steam Boilers, ought 
to be carefully studied, and always at hand."— Coll. Guardian. 

" The book is vkry useful, especially to steam users, artisans, and "young Engineers."— 

By the SAMS Author, 


they Occur, and How to Prevent their Occurrence. A Practical Hand- 
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Price 3s. 



Just Out. In Crown $vo, Handsome Cloth, With Numerous 

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A Text-Book of Workshop Practice in General Tool Grinding, 

and the Design, Construction, and Application 

of the Machines Employed. 


R. B. HODGSON, A.M. Inst. Mech.E., 

Author of "Machines and Tools Employed in the Working of Sheet Metals." 

Introduction. — Tool Grinding. — Emery Wheels. — Mounting Emery Wheels. 
— Emery Rings and Cylinders. — Conditions to Ensure Efficient Working. — 
Leading Types of Machines. — Concave and Convex Grinding. — Cup and Cone 
Machines. — Multiple Grinding. — " Guest " |Universal and Cutter Grinding 
Machines. — Ward Universal Cutter Grinder. — Press. — Tool Grinding. — Lathe 
Centre Grinder. — Polishing. — Index. 

"Deals practically with every phase of his subject." — Ironmonger. 

Fifth Edition. Folio, strongly half-bound, 21/. 


Computed to Four Plaees of Decimals for every Minute 

of Angle up to 100 of Distance. 

For the use of Surveyors and Engineers. 


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* # * Published with the Concurrence of the Surveyors- General for New South 

Wales and Victoria. 

"These who have experience in exact Survby-wo&k will best know how to appreciate 
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ensure to every user, and as every Surveyor in active practice has felt the want of such 
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Formerly Professor of Electrical Engineering, The Glasgow and West of Scotlana 

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In Large Crown %vo. Fully Illustrated. 


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Book on). For Advanced and " Honours " Students. By Prof. Jamieson, 
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VoL I. — Comprising Parti.: The Principle of Work and its applica- 
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'The work has very high qualities, which may be condensed into the one word 
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In Preparation. 300 pages. Crown Svo. Profusely Illustrated. 

Modern Electric Tramway Traction: 

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For the Use of Electricians and Engineers. Pocket Size. Leather, 
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Urtt *•#/«# Pr9f§9§or •/ Olvll EnglMtrtng In th§ Untutrafty §f Glasgow. 


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let* 9§§rttary U tha InrtttuU •/ Engtnaar* and 8hl/ttulkhr$ In f tefttamt 


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• * 


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Second Edition, Revised. With Diagrams. Price 2a. 

Latitude and Longitude: 

How to Find them. 

By W. J. MILLAR, C.E., 

LaU Soerotarf to tho Inst, of Engineers and Shipbuilders in Scotland. 

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_ M I. _ | - 

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Shifting Cargoes, 
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ftc, Ac. 

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Steel Ships: Their Construction and Maintenance. 

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Fourteenth Bdltlon, Revised, Prlee 21a. 

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By A. E. S E A T N, M. Inst. C. E., M. Inst. Meeh. B.. 


General Contents. — Part I. — Principles of Marine Propulsion. 
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M.I.M.E., M.LE1.E., HLMinJE., Whit Sch., M.Ord.MeiJL 



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Showing at a glance the Mutual Conversion of Measurements 

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An Introduction to the Study of Force and Motion. 

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A Manual for Users of Steam and Water. 

9n P*or. FRANZ SCHWACKH0FER of Vienna, and 

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A Guide to the Construction of Works for the Prevention of the 
Pollution by Sewage of Rivers and Estuarie* 

By W. SANTO CRIMP, M.Inst.C.E., F.G.S., 

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